Mining  Dent* 


COAL 
MINING   COSTS 


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COAL 
MINING   COSTS 


BY 

A.  T.  SHURICK 

•  ? 

MEMBER  AMERICAN  INSTITUTE  OF  MINING  ENGINEERS, 

FORMERLY   ASSOCIATE   EDITOR    "COAL  AGE,"    ETC. 


FIRST  EDITION 


McGRAW-HILL  BOOK  COMPANY,  INC. 

NEW  YORK:   370  SEVENTH  AVENUE 

LONDON:  6  &  8  BOUVERIE  ST.,  E.  C.  4 

1922 


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COPYRIGHT,  1922,  BY 
A.  T.  SHURICK 


PREFACE 

THERE  are  books  on  costs  in  all  the  important  branches  of 
engineering  except  coal  mining.  The  author  has  waited  patiently 
the  advent  of  a  similar  work  in  his  chosen  field  and  none  having 
been  forthcoming  he  has  made  bold  to  venture  the  effort  himself. 

Obviously  a  work  of  this  character  cannot  be  up-to-date  as 
to  costs  in  dollars  and  cents  because  of  the  wide  fluctuations  in 
the  purchasing  power  of  the  dollar  in  labor,  equipment  and 
material,  particularly  during  the  last  five  years.  There  has 
been  no  hesitancy,  therefore,  in  using  data  of  a  number  of  years 
back  so  long  as  the  subject  discussed  is  still  in  general  use  as 
in  the  case  of  the  comparative  costs  of  wood  and  masonry  brat- 
tices, etc.  The  reader  will  have  no  difficulty  in  interpolating 
the  figures  given  to  conform  to  current  standards  and  to  facili- 
tate this  a  table  giving  all  the  wage  scales  in  the  Central  Com- 
petitive District  since  1898  has  been  given  on  page  158.  Thus 
if  a  certain  piece  of  work  required  the  services  of  three  men 
for  a  certain  number  of  days,  say  a  decade  ago,  it  is  a  relatively 
simple  matter  to  estimate  the  cost  in  terms  of  prevailing  wage 
scales.  Care  has  been  exercised  to  give  the  year  during  which 
the  different  examples  cited  occurred  in  order  that  this  inter- 
polation can  readily  be  effected. 

A  great  deal  of  valuable  data  of  an  abstract  nature  has  been 
obtained  from  the  various  State  and  Federal  government  reports 
but  in  general  it  has  been  the  endeavor  to  hold  more  to  specific 
costs.  Thus  the  cost  of  haulage  or  the  cost  of  doing  a  piece 
of  work  under  certain  well  defined  conditions  has  been  accepted 
as  of  more  value  than  the  average  cost  of  mining  for  a  certain 
district  or  the  capital  investment  per  ton  of  capacity,  etc.  In 
other  words  it  has  been  the  aim  to  make  the  work  essentially 
practical. 

Only  a  few  of  the  best  systems  of  mining  have  been  discussed 
but  these,  it  is  believed,  have  been  covered  in  greater  detail  than 
in  works  dealing  with  this  subject  alone.  The  reason  for  this 


48265,; 


vi  PREFACE 

is  that  unless  full  particulars  concerning  all  phases  of  working 
under  any  system  are  made  clear  beyond  all  peradventure  the 
cost  figures  are  worthless. 

To  insure  a  thorough  treatment  of  all  subjects  taken  up  it 
was  deemed  advisable  to  limit  the  present  volume  to  under- 
ground costs  alone.  A  great  deal  of  valuable  data  on  outside 
costs  has  been  assembled  in  the  course  of  the  present  work  and 
it  is  thought  the  publication  of  this  in  a  separate  volume  at  a 
later  date  will  enhance  the  value  of  the  completed  work  more 
than  if  an  attempt  were  made  to  straddle  the  two  fields  in  the 
present  volume. 

It  is  thought,  if  anything,  the  book  errs  on  the  side  of  con- 
servatism ;  old  and  tried  methods  only  have  been  accepted,  newer 
systems  and  appliances,  some  of  which  are  very  promising  at 
this  time,  having  been  used  with  caution.  As  an  example  the 
use  of  the  underground  loading  machine  or  the  combined  mining 
and  loading  machine  has  not  become  sufficiently  general  as  yet 
to  justify  the  amount  of  space  devoted  to  say,  the  mine  motor 
which  is  now  an  accepted  part  of  nearly  every  mining  operation. 

In  concluding,  the  author  wishes  to  make  sincere  acknowledg- 
ment to  the  many  friends  who  have  furnished  him  with  much 
valuable  material  and  still  more  valuable  suggestions  and  advice ; 
to  the  various  engineering  societies,  the  papers  of  which  have 
been  quoted  so  freely  throughout  the  book ;  to  the  large  engineer- 
ing companies,  several  of  which  in  particular  have  been  untiring 
in  their  efforts  to  furnish  certain  special  requests  for  material; 
and  to  the  technical  press  which  has  been  drawn  upon  liberally. 

A.  T.  SHURICK. 

DEARBORN,  MICH. 
January,  1922. 


CONTENTS 

SECTION  I 

MINING  COSTS 

PAGE 

Government  statistical  data  regarding  costs — Method  of  computing  tax 
returns — Comparison  of  costs  and  distribution  of  revenue  for  German 
and  Middle  Western  Mines — Conditions  where  operations  may  be 
conducted  at  an  apparent  loss — Systems  of  mining — Methods  of 
working  in  the  Pocahontas  field — Connellsville  systems — Compara- 
tive costs  of  mining  different  thicknesses  of  coal — Conveyor  system — 
Mining  machinery — Cutting  machines — Installation  and  operating 
costs  —  Comparative  cost  for  alternating  and  direct  current  for 
machines — Arc  wall  cutters — Post  punchers — Repair  costs — Machine 
bits — Loading  machines — Mining  and  loading  machines — Blasting — 
Dynamite — Hydraulic  cartridges — Shot  firing — Daymen — Miner's 
wages — Losses  from  idle  time — Economic  aspects  of  conservation — 
Use  of  longwall  to  effect  conservation 1 


SECTION  II 
SHAFT  SINKING 

Operations — Circular  or  rectangular — Equipment — Sinking  costs — Shaft 

linings — Rates  of  progress — Reports  and  contract  forms 176 

SECTION  III 
HAULAGE   COSTS 

Tractive  effort,  drawbar  pull  and  rating  of  mine  motors — Number  and 
size  of  motors  required — Gathering  locomotives — Costs — Motor  losses 
— Power  costs — Line  costs  and  losses — Bonding — Computing  losses 
in  bonds — Track  costs — Haulage  grading  estimates — Haulage  track 
curves — Rails — Track  frogs — Track — Mine  cars — Rope  haulage — 
Gravity  planes — Endless  rope  haulage — Wire  rope  lubrication — Com- 
parative costs  of  different  systems  of  haulage — Comparison  of  all 
systems — Animal,  compressed  air  and  electric  haulage  costs — 
Single-  and  two-stage  air  motors — Compressed-air  and  animal  haul- 
age costs — Electric  motor  and  animal  haulage  costs — Costs  and  care 
of  mules — Gasoline  motor  vs.  animal  haulage — Storage  battery  and 
trolley  motors  haulage  costs — Compressed-air  and  electric  haulage 
costs 219 

vii 


viii  CONTENTS 

PAGB 

SECTION  IV 
TIMBERING  COSTS 

Ho     to  buy  timber — Timber  used  and  costs  for  the  United  States — Com- 
puting size  of  timber — Timber  framing  equipment — Timber  pre- 
61  vatives — Steel  timbering — Concrete  timber — Cement   gun — Re- 
aiming  timbers 365 

SECTION  V 
MISCELLANEOUS  INSIDE   COSTS 

nneling  costs — American  and  foreign  tunneling  records  compared — Cost 
of  rock  tunnel  at  a  coal  operation — Some  examples  of  tunnel  costs — 
Drill  steel — Explosives — Ventilating  costs — Power  required — Speci- 
fication of  fans — Change  in  volume  of  air  required — Mine  lighting — 
Portable  electric  lamps — Oil  and  acetylene  lamps — Oil  lamps — Cost 
of  undergrounds  tables — Overcasts — Stoppings  and  overcasts — Com- 
parison of  doors  and  overcasts — Cost  of  stone  and  wood  brattices — 
Refuge  chambers — Mine  sprinkling  costs 406 


GOAL  MINING  COSTS 


SECTION  I 
MINING  COSTS 

The  U.  S.  Census  report  for  the  year  of  1909  shows  that  the 
value  of  the  Pennsylvania  anthracite  produced  that  year  was 
$148,957,894.  The  total  gross  expenses  amounted  to  $139,110,- 
444,  from  which  should  be  deducted  $4,864,844  made  from 
charges  to  miners  for  explosives,  oil  and  blacksmithing,  making 
the  net  expenses  $134,245,600. 

The  gross  expenses  are  itemized  as  follows : 

Services: 

Salaries $  4,572,489 

Wages 92,169,906    $  96,742,395 

Supplies : 

Fuel  and  power 3,189,279 

Other  supplies 23,472,809        26,662,088 

Royalties 7,969,785 

Miscellaneous 7,736,176 


Total  gross  expenses $139,110,444 

Deductions 4,864,844 


Net  expenses $134,245,600 

The  total  production  in  1909  amounted  to  72,215,273  long 
tons,  so  that  the  average  value  per  ton  for  the  output  in  that 
year  was  $2.06 ;  the  average  cost  per  ton  was  $1.86 ;  and  the  net 
returns  on  the  operations  for  the  year  were  $14,712,294,  or  an 
average  of  20c  per  ton.  This  at  first  glance  looks  like  a  fair 
return,  but  attention  must  be  called  to  the  fact  that  the  Census 


2  COAL  MINING  COSTS 

figures  of  cost  make  no  allowance  for  interest  on  capital  invested 
or  borrowed,  and  no  offsetting  charges  for  amortization  or  depre- 
ciation. 

According  to  the  returns  to  the  Bureau  of  the  Census,  the 
entire  capital  invested  in  anthracite  mining  in  1909  was  $246,- 
700,000,  which  may  appear  rather  inadequate  when  one  considers 
the  magnitude  of  the  industry,  and  an  annual  production  of 
$150,000,000  (in  1911  the  output  was  valued  at  $175,189,392 
and  in  1912  it  was  $177,622,626),  but  these  are  the  figures 
reported  by  the  Census  Bureau.  If  on  this  capitalization  an 
allowance  of  4  per  cent  be  made  for  interest,  the  net  returns 
for  the  year  amounted  in  round  numbers  to  $4,844,000. 

If  new  breakers  and  other  equipment  are  charged  into  operat- 
ing expenses  no  allowance  need  be  made  for  depreciation,  but 
the  exhaustion  of  from  75,000,000  to  80,000,000  tons  from  the 
reserves  every  year  should  have  some  amortization  charged 
against  it  and  if  5c.  a  ton  be  allowed  the  margin  of  $4,800,000 
is  practically  wiped  out. 

The  figures  covering  the  cost  and  value  of  bituminous  coal 
show  even  more  striking  comparisons.  There  are  some  slight 
differences  in  the  statistics  of  production  between  the  Census 
figures  and  those  published  by  the  United  States  Geological 
Survey  for  the  reason  that  the  Census  investigations  excluded 
mines  having  a  production  of  less  than  1000  tons,  whereas  the 
Survey  includes  every  small  country  bank  from  which  it  can 
secure  a  report.  For  1909  the  Survey  showed  a  bituminous  coal 
production  of  379,744,257  short  tons  valued  at  $405,486,777, 
and  the  Census  Bureau  showed  a  production  of  376,865,510  tons 
valued  at  $401,577,477,  the  difference  being  about  3,000,000  tons 
in  quantity  and  $4,000,000  in  value — less  than  1  per  cent  in 
either  case.  As  the  Census  figures  for  cost  of  mining  are  the 
basis  of  this  discussion,  the  Census  figure  of  production  is  also 
used. 

The  total  value  of  the  bituminous  production,  as  already 
stated,  was  $401,577,477,  and  the  mining  expense  of  producing 
this  value,  including  salaries  of  officers,  was  $378,159,282.  As 
in  the  case  of  anthracite,  the  expenses  of  production  do  not 
include  any  charges  for  depreciation,  amortization,  or  interest 
on  capital  invested  or  borrowed.  The  expenses  are  divided  as 
follows : 


MINING  COSTS 

Salaries $  20,417,392 

Wages 282,378,886 

Supplies 45,345,932 

Royalties • .  / 12,035,900 

Miscellaneous 17,961,172 


Total $378,159,282 

From  this  it  appears  that  75  per  cent  of  the  total  cost  and 
70  per  cent  of  the  total  value  was  spent  in  wages.  Salaried 
officials  got  less  than  5.5  per  cent. 

The  total  capital  invested  in  the  bituminous  coal  mines  of 
the  United  States  in  1909  was,  according  to  the  Census  bulletin, 
in  round  numbers  $960,000,000  ($960,289,465),  and  this  does  not 
appear  as  if  there  were  very  much  over- valuation,  whatever  the 
capitalization  may  be  as  represented  by  stock  issue.  The  dif- 
ference between  the  value  of  the  product  and  the  expense  of 
producing  it  was  $23,440,000  in  round  numbers  or  a  fraction 
over  2.5  per  cent  on  the  capital. 

According  to  the  figures  compiled  by  the  Bureau  of  the 
Census,  the  amount  paid  in  wages  was,  in  1909,  above  80  per 
cent  of  the  total  selling  value  of  coal  at  the  mine  mouth.  From 

1909  to  1913  there  were  two  wage  increases  granted — one  in 

1910  of  5.55  per  cent  and  another  in  1912  of  5.26  per  cent. 
These  increases  brought  the  wage  cost  per  ton  of  coal  produced 
to  92.44c.  in  1913. 

In  1913  the  average  selling  price  of  coal  at  the  mines  in 
Illinois  was  $1.14  and  in  Indiana  $1.11  per  ton.  This  leaves  a 
margin  of  only  21. 6c.  in  Illinois  and  18. 6c.  in  Indiana.  Out 
of  this  must  be  paid  the  cost  of  material  used  at  the  mines;  the 
cost  of  making  sales;  all  officers'  salaries;  general  expenses; 
insurance  (liability,  fire,  storm,  etc.)  ;  taxes  (including  tax  on 
plant  and  mineral  rights)  ;  interest  on  the  investment;  deprecia- 
tion of  plant;  royalties  or  charges  for  the  exhaustion  of  coal. 

The  report  of  the  Bureau  of  the  Census  for  1909  showed 
that,  without  allowing  for  any  interest  charge  on  the  investment 
or  for  amortization  of  property,  the  so-called  net  returns  in 
Illinois  and  Indiana  were  only  3c.  per  ton  in  Illinois  and  less 
than  Ic.  per  ton  in  Indiana. 

The  average  royalty  paid,  however,  in  these  two  states  on 
coal  recovered  under  lease  is  5c.  per  ton,  and  the  average  present 
valuation  of  coal  land  is  such  as  to  require  a  minimum  amor- 


COAL  MINING  COSTS 


tization  charge  of  3c.  per  ton  to  recover  such,  land  value  within 
the  period  of  the  mine's  life.  It  will  therefore  be  seen  that  in 
even  so  good  a  year  as  1913  an  actual  profit  return  was  impos- 
sible. As  existing  facts  show,  the  industry  sustained  a  sub- 
stantial deficit  in  these  two  states. 

The  average  value  per  ton  of  all  the  bituminous  coal  produced 
in  the  United  States  was  $1.07  and  the  costs  averaged  a  fraction 
of  a  cent  over  $1,  so  that  the  margin  of  profit  to  cover  interest, 
depreciation  and  amortization  was  a  little  less  than  7c.  a  ton. 
In  some  states  the  expenses  exceeded  the  returns.  Take  Arkansas, 
for  instance,  where  the  expenses  totaled  $3,630,526  and  the  value 
of  the  product  was  $3,508,590.  Other  instances  were : 


Value  of  Product 

Expenses 

Iowa  

$12,682,106 

$12,816,076 

Kentucky                             

9,940,485 

10,127,987 

Tennessee 

6  548  515 

6  691  482 

Oklahoma                 

6,185,078 

6,536,441 

Virginia 

4,336,185 

4  392,440 

Pennsylvania,  by  long  odds  the  most  important  producer, 
with  an  output  of  137,300,000  tons,  showed  a  total  of  expenses 
of  $117,440,000  and  of  value  of  $129,550,000  a  balance  on  the 
profit  side  of  a  little  over  $12,000,000,  or  about  3l/3  per  cent  on 
the  capital  invested,  $358,600,000.  The  four  competitive  states, 
West  Virginia,  Illinois,  Ohio  and  Indiana,  which  rank  second, 
third,  four  and  fifth,  respectively,  in  producing  importance,  all 
show  such  narrow  margins  between  income  and  outlay  that 
profits  are  infinitesimal.  The  figures  follow: 


Value  of  Product 

Expenses 

Difference 

West  Virginia  
Illinois                 .... 

$  44,344,067 
53,030,545 

$  43,024,716 
51,697,504 

$1,319,351 
1,333,041 

Ohio 

27  353  663 

27  153  497 

200,  166 

Indiana    

15,018,123 

14,906,831 

111,292 

$139,746,398 

$136,782,548 

$2,963,850 

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6  COAL  MINING  COSTS 

These  four  states  with  an  aggregate  production  of  a  little 
more  than  the  bituminous  output  of  Pennsylvania,  showed  a 
total  of  less  than  $3,000,000  as  the  excess  of  receipts  over 
expenses.  The  capital  invested  in  the  coal-mining  industry  in 
these  states  was  something  over  $310,000,000,  so  that  the  returns 
on  the  capital  were  less  than  1  per  cent. 

The  United  States  Fuel  Administration,  organized  during  the 
war  emergency,  was  empowered  to  exact  the  most  intimate  details 
as  to  operating  costs  and  profits  in  the  coal  industry  under 
severe  penalities  for  omissions  or  incorrect  returns.  The  En- 
gineers Committee  of  the  Fuel  Administration,  headed  by  some 
of  our  most  prominent  engineers  and  equipped  with  an  excellent 
organization  for  assembling  and  correlating  this  mass  of  material 
compiled  a  report  on  general  production  costs  unexcelled  in  the 
history  of  the  industry  for  its  authenticity  and  accuracy. 

The  accompanying  table  is  a  summary  showing  reported  and 
adjusted  costs,  prices  fixed  and  tonnage  for  all  the  principal 
districts  to  August  12,  1918.  The  diagram  Fig.  1  shows  these, 
in  general,  before  the  labor  increase  of  November,  1917,  com- 
pensated for  by  the  45c.  general  advance  in  coal  prices,  giving 
the  average  costs,  "bulk  lines,"  and  prices  fixed  for  practically 
all  districts  in  the  country  as  of  August  and  September,  1917, 
and  covers  about  95  per  cent  of  the  total  output  of  bituminous 
coal  for  the  period  stated. 

The  costs  for  each  district  in  the  proportion  of  its  output 
to  the  total  tonnage  studied  are  shown  in  heavy  lines,  the  * '  bulk 
lines"  are  shown  by  medium  lines,  and  the  prices  fixed  are 
indicated  by  light  lines.  The  diagram  also  shows  the  weighted 
average  costs,  "bulk  lines,"  and  prices  fixed  for  the  tonnage 
included,  and  effectively  disposes  of  the  widely  circulated  asper- 
sions of  profiteering,  of  which  the  industry  was  so  freely  accused 
by  people  having  no  knowledge  of  the  facts  or  willfully  misrep- 
resenting them. 

Diagram  Fig.  2  shows  the  same  data  for  the  principal  dis- 
tricts for  the  full  year  1918,  giving,  however,  only  reported 
costs  and  prices  fixed,  the  prices  having  been  fixed  on  the  August- 
September,  1917,  data,  and  changed  only  by  the  45c.  allowed 
November  1,  1917,  to  compensate  for  the  labor  increase  of  that 
date,  reduced  May  24,  1918,  to  35c.  in  consideration  of  equal 
car  distribution  ordered  at  that  time. 


MINING  COSTS 


The  weighted  average  margin  between  costs  and  prices  for 
practically  the  entire  bituminous  coal  production  of  the  country 
was  but  45.  6c.,  and  between  the  "bulk  line,"  which  represents 


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Administrator,  was  but  26c. 

As  it  is  known  that  the  capital  invested  per  ton  of  yearly 
output  in  bituminous  mines  ranges  from  $2  to  nearly  $8,  interest 
on  which  is  included  with  other  mines  in  the  "margin"  not 
included  in  the  charted  costs,  it  is  evident  that  taking  the 


COAL  MINING  COSTS 


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13 


industry  as  a  whole  no  excessive  price  allowances  were  given. 
If  prices  had  been  fixed  at  a  point  high  enough  to  even  cover 
the  highest  costs  reported  in  each  district,  the  result  would  have 
been  to  add  over  a  billion  dollars  to  the  price  paid  for  coal, 


with  probable  labor  disturbances,  in  an  effort  to  obtain  some  of 
the  abnormal  profits  which  would  have  gone  to  the  great  majority 
of  the  tonnage,  so  serious  as  to  have  probably  decreased  rather 
than  increased  the  tonnage,  which  was  in  fact  ample  for  all 
the  needs  of  the  country. 


14  COAL  MINING  COSTS 

The  prices  fixed  from  this  complete  investigation  of  costs 
have  shown  in  many  cases  a  remarkable  compliance  with  eco- 
nomic laws.  For  instance,  in  Illinois  the  cost  of  coal  from  the 
different  price  districts  delivered  in  Chicago  was  found  to  be 
practically  identical,  showing  that  the  mining  of  the  higher- 
cost  coal  is  due  to  its  proximity  to  the  principal  market  and 
the  lower  resulting  transportation  costs.  High-grade  coal  shipped 
by  lake  and  rail  to  Minneapolis  was  found  to  cost  precisely  the 
same  per  heat  unit  as  a  lower-grade  coal  shipped  a  much  less 
distance  all  rail. 

Anthracite  prices  as  fixed  by  the  President  Aug.  23,  1917, 
with  an  adjustment  for  the  labor  increase  of  Dec.  1,  1917, 
were  the  subject  of  an  intensive  study  by  the  committee  im- 
mediately after  the  first  charting  of  bituminous  costs  was  com- 
pleted. 

A  technical  paper  giving  the  methods  adopted  and  the  results 
of  this  analysis  was  presented  before  the  American  Institute  of 
Mining  Engineers,  February,  1919,  and  the  following  is 
abstracted  from  this  paper : 

The  adjustments  of  cost  from  a  reported  to  a  price-fixing  basis,  as 
described  for  the  bituminous  methods,  were  applied  but  showed  only  minor 
adjustments  as  necessary. 

The  great  spread  in  anthracite  prices  on  the  varying  sizes,  which  for 
the  6 -month  period  under  review  ranged  in  average  from  $5.244  for  nut 
to  $2.074  for  barley  coal,  makes  the  question  of  the  percentage  of  sizes 
produced  at  the  different  collieries  a  vital  one.  The  realization  with  the 
same  prices  for  each  size  must  be  within  very  wide  limits,  when  it  is 
considered  that  the  percentage  of  prepared  coal  reported  from  different 
collieries  varied  from  over  80  per  cent  to  below  30  per  cent  for  fresh-mined 
coal.  Hence,  as  the  spread  in  prices  for  the  various  sizes  must  be  predi- 
cated on  some  percentage,  it  is  essential  to  find  some  method  of  adjustment 
to  allow  for  this  variation.  The  logical  method  of  adjustment  is  to 
calculate  actual  costs  to  costs  as  of  the  standard  percentage  of  sizes,  so 
that  the  margin  between  the  adjusted  costs  and  the  average  realization 
shall  be  the  actual  margin  for  each  colliery  between  its  actual  costs  and 
actual  realization  due  to  its  particular  percentage  of  sizes.  As  a  base  for 
realization  the  actual  percentage  of  sizes  for  fresh-mined  coal  for  the 
6  month  period  was  adopted.  This  percentage  is  given  in  the  following 
table. 


MINING  COSTS 


15 


PERCENTAGE  OF  SIZE  OF  FRESH-MINED  COAL 


Size  of  Coal 

MESH,  IN  INCHES 

PERCENTAGE  OF  SIZES 

Through 

Over 

Fresh- 
mined 

Washery 

Fresh- 
mined  and 
Washery 

Round 

Round 

Broken  

4| 
3f      31 
2^     21 

if     14 
f 
4 
A     1 
&     A 
A     1 
A     A 

3|      31 
2A     21 

if     14 
f 
1 
A     1 
A     A 
A     A 
A     A 

6.8 
14.6 
19.6 
24.7 
9.1 
11.6 
3.2 
4.9 
3.9 
1.6 

0.4 
1.2 
2.3 
10.1 
10.0 
21.4 
14.9 
27.5 
8.8 
3.4 

6.2 
13.5 
18.2 
23.5 
9.2 
12.4 
4.2 
6.8 
4.3 
1.7 

Eee 

Stove 

Nut      

Pea                

Buckwheat 

Rice           

Barley 

Boiler  

Screenings  

For  adjustment  as  a  base  for  fixing  a  spread  of  prices  the  percentages 
used  were,  taken  at  even  figures,  prepared  65  per  cent,  pea  9  per  cent, 
buckwheat  12  per  cent,  and  smaller  14  per  cent. 

The  adjustment  finally  arrived  at  after  long  study  was  tested  on  actual 
reports  from  collieries  having  percentages  that  varied  from  over  80  per 
cent  to  under  30  per  cent  prepared  coal  and  was  found  to  be  correct  within 
a  maximum  variation  of  less  than  1%  per  cent.  It  was  as  follows: 


For  Each  1  Per  Cent  Variation 

Above  Standard, 
Per  Cent 
Deduction 

Below  Standard, 
Per  Cent 
Addition 

Prepared  

1.20 

1.20 

Pea                                      .              

0.85 

0.85 

Buckwheat 

0  75 

0.75 

Smaller  

0.50 

0.50 

As  examples  of  the  working  of  this  adjustment  with  prices  assumed 
at  about  the  average  for  the  6  months  and  taking  mines  well  away  from 
average  percentage  of  sizes,  the  following  may  be  cited: 


16 


COAL  MINING  COSTS 


Size 

Base 

Sizes, 
Per 
Cent 

Base 
Price 

Reali- 
zation 

Mine 
A 
Sizes, 
Per 
Cent 

Correc- 
tion, 
Per 
Cent 

Actual 
Reali- 
zation 

Mine 
B 
Sizes, 
Per 
Cent 

Correc- 
tion, 
Per 
Cent 

Reali- 
zation 

Prepared    

65 
9 
12 
14 

$5.10 
3.70 
3.20 
2.20 

$3.315 
0.333 
0.384 
0.308 

73.1 
6.4 
10.4 
10.1 

-9.72 
+  2.21 
+  1.20 
+  1.95 

$3.730 
0.237 
0.333 
0.222 

55.1 
15.3 
13.7 
15.9 

+  11.880 
-   5.355 
-    1.275 
-    0.950 

$2.810 
0.566 
0.438 
0.350 

Pea              

Buckwheat 

Smaller             

Total  

100 

$4.340 

100.0 

-4.36 

$4.522 

100.0 

+  4.30 

$4.164 

Assume  cost  for  each  mine $4 . 000 

Actual  margin 0 . 522 


$4.000 
0.164 


Standard  realization $4 . 340 

Calculated  cost  as  of  standard  per  cent  sizes, 

$4X0.9564%=        3.826  $4X104.30% 


Calculated  margin $0. 514 


$4.340 
4.172 
$0.168 

The  correction  for  mine  A  is  then  — 4.36  per  cent  and  the  adjusted 
cost  $3.826,  showing  51.4c.  margin  on  the  $4.34  standard  realization  against 
52. 2c.  actual  margin.  Similarly,  for  mine  B,  the  correction  is  -{-4.30  per 
cent,  giving  an  adjusted  cost  of  $4.172  and  a  margin  of  16. 8c.  as  compared 
with  the  actual  margin  of  16. 4c.  Thus  the  adjusted  costs  on  the  chart 
bear  a  true  relation  to  the  realization  received  from  a  scale  of  prices  for 
the  various  sizes  based  on  the  standard  or  average  percentage  of  sizes 
adopted  as  a  base,  regardless  of  the  actual  percentage  of  sizes  produced 
by  each  operation,  and  prices  can  be  fixed  from  the  chart  line  of  adjusted 
costs  which  will  result  in  giving  each  mine  its  intended  margin.  The 
correction,  of  course,  is  an  allocation  based  on  realization  from  the  different 
sizes  and  could  be  made  more  accurately  by  taking  into  account  each  size 
produced,  but  at  the  cost  of  more  time  than  was  available  for  the  work. 
With  a  material  variation  in  price,  different  factors  of  correction  should  be 
calculated. 

A  large  percentage  of  the  anthracite  coal  is  owned  in  fee  by  operators, 
who  also  lease  tracts  contiguous  to  their  fee  holdings.  As  all  report  royal- 
ties on  the  basis  of  tonnage  produced,  the  general  average  15. 5c.  per  ton 
reported  is  misleading.  The  actual  average  royalty  reported  by  operators 
mining  generally  from  leased  lands  was  33.25c.  and  by  those  generally 
mining  from  fee  lands,  5.5c.  As  relatively  few  operators  mine  exclusively 
from  either  class  of  lands,  no  data  are  available  to  show  the  actual  average 
royalties  paid,  but  it  is  believed  that  the  present  average  would  be  approxi- 
mately 40c.  per  ton. 

A  few  leases,  notably  those  made  by  the  trustees  of  the  Girard  estate, 
owned  by  the  city  of  Philadelphia,  base  the  royalty  payments  on  a  per- 
centage of  the  sale  price  of  the  coal  at  the  mines  instead  of  requiring 
fixed  royalties.  This  percentage  varies  from  15  per  cent  to  as  high  as  28 
per  cent  of  the  price.  As  the  labor  war  bonuses  materially  add  to  the  sale 


MINING  COSTS 


17 


price,  these  have  resulted  in  excessive  royalties  and  serious  embarrassment 
to  the  operators,  who  were  not  allowed  to  increase  the  price  of  coal  suffi- 
ciently to  even  fully  absorb  this  additional  labor  cost  and  by  whom  the 
extra  royalties  must  be  paid  out  of  already  narrow  margins. 

Cost  charts  were  made  from  averages  of  the  6  months,  showing  both 
the  reported  and  the  adjusted  costs  for  standard  fresh-mined  white  ash 
anthracite,  both  by  collieries  and  by  operating  companies.  As,  in  the 
prices  fixed  by  the  President  Aug.  23,  1917,  a  differential  of  75c.  per  ton 
on  pea  size  and  above,  equivalent  to  52.95c.  per  ton  on  all  sizes,  was 
established  for  the  independent  operators  over  certain  companies  with  rail- 
road affiliation,  generally  known  as  the  l '  companies. ' ' 

AVERAGE  AND  BULK  LINE  COSTS  OF  WHITE  ASH  COAL 


Description 

Costs, 
Averages 
Returned 

Costs, 
Adjusted 

Cost, 
90  Per  Cent 
Bulk  Line 

Excluding  washery  coal: 
All  operations,  each  colliery  separate  
All  company  operations,  each  colliery  separate  
All  independent  operations,  each  colliery  separate  .  .  . 
All  operations,  each  company  operating  two  or  more 

$3.85 
3.71 
4.37 

3.85 

$3.91 
3.79 
4.36 

3.91 

$4.80 
4.65 
4.97 

4  38 

Including  washery  coal: 
All  operations,  each  company  operating  two  or  more 
collieries  consolidated  

3.57 

3.77 

4.36 

AVERAGE  PRICES  RECEIVED  FOR  WHITE  ASH  COAL 


FRESH-MINED 

BANK 

TOTAL,  INCLUDING 

COAL 

COAL 

BANKS 

Size 

Per 

Average 

Per 

Average 

Per 

Average 

Cent 

Price 

Cent 

Price 

Cent 

Price 

6  8 

$4  889 

0.4 

$4.416 

6.2 

$4  886 

Eee 

14  6 

5  028 

1  2 

4.815 

13.5 

5  027 

Stove 

19  6 

5  161 

2.3 

5.060 

18.2 

5.160 

Nut 

24  7 

5  244 

10.1 

5.246 

23.5 

5.244 

Pea 

9   1 

3  687 

10.0 

3.696 

9.2 

3.698 

prepared  and  pea  

74.8 

$4.959 

24.0 

$4.544 

70.6 

$4.947 

Buckwheat  

11.6 

$3.342 

21.4 

$3.213 

12.4 

$3.324 

Rice  

3.2 

2.482 

14.9 

2.452 

4.2 

2.473 

Barley  

4.9 

2.231 

27.5 

1.767 

6.8 

2.074 

Boiler  

3.9 

2.341 

8.8 

2.123 

4.3 

2.304 

Screenings  

1.6 

2.202 

3.4 

1.555 

1.7 

2.162 

small  sizes  

25.2 

$2  .  795 

76.0 

$2.339 

29.4 

$2.697 

Grand  total  

100.0 

$4.414 

100.0 

$2.868 

100.0 

$4  .  285 

18 


COAL  MINING  COSTS 


The  prices  received  by  the  companies  and  independents  have  not  been 
separately  averaged,  but,  calculating  on  the  differential  and  assuming  the 
percentages  the  same  for  companies  and  independents  which  is  only  approxi- 
mately the  case,  the  selling  price  of  fresh-mined  coal  would  average  for 
companies  $4.287,  and  for  independents,  $4.817.  Margins  over  reported 


costs  of  companies  would  be  58c.,  and  for  independents  45c.,  with  a  general 
average  margin  for  all  fresh-mined  coal  of  56c.,  and  for  all  coal,  including 
washery,  of  71c.  per  ton,  and  under  "bulk  line"  costs,  fresh-mined  com- 
panies, 36c.;  independents,  15c. ;  total,  39c.,  including  washeries  consolidated 
sheets  total  of  7.5c. 

These  margins  include  all  expenditures  for  Federal  income  and  excess- 


MINING  COSTS 


19 


profit  taxes,  selling  expenses,  interest  charges,  expenditures  for  improve- 
ments and  developments  to  increase  output,  excess  of  capital  expenditures 
over  normal  cost,  and  all  profit  on  the  investment  of  about  $8  per  ton 
annual  output. 

Effective  Dec.  1,  1917,  a  labor  war  bonus,  ranging  from  60c.  to  $1.10 
per  day  for  labor  and  25  per  cent  for  contract  miners  was  granted  over 
and  above  the  wage  scales  effective  by  agreement  Apr.  1,  1916,  expiring 
Apr.  1,  1920,  and  the  prices  fixed  Aug.  23,  1917,  and  modified  Oct.  1, 
1917,  by  reducing  pea  coal  60c.  per  ton,  were  increased  by  35c.  per  ton 
to  compensate  for  this  labor  increase.  The  actual  reported  increase  in 
labor  costs  due  to  this  advance  was  figured  by  the  Federal  Trade  Commission 
from  the  operators'  reports  to  be  60. 3c.  From  the  actual  pay-roll  figures 
later  obtained  by  the  United  States  Fuel  Administration,  this  increase 
was  found  to  be  76. 3c.  per  ton. 

Effective  Nov.  1,  1918,  a  second  labor  war  bonus  was  granted.  The 
calculated  increase  in  cost  due  to  this  is  shown  in  Fig.  3,  on  which  the 
increases  for  each  operator  are  found  by  figuring  from  the  pay  rolls  for 
the  6  months  the  actual  increase  in  pay  which  would  have  been  given, 
applying  the  Nov.  1,  1918,  increases,  and  dividing  by  the  6  months'  tonnage 
of  the  colliery.  This  line,  adjusted  to  per  cent  of  sizes,  and  compared 
with  the  adjusted  cost,  shows  an  increase  in  cost  of  74. Ic.  As  this  was 
necessarily  applied  to  the  prepared  and  pea  sizes,  70.6  per  cent  of  the 
total,  the  increase  on  those  sizes  was  $1.05  per  ton,  which  increase  was 
allowed  to  balance  the  increased  cost  of  labor. 

Except  for  the  two  increases  to  compensate  for  labor  increases  just 
noted  and  the  reduction  Oct.  1,  1917,  of  the  pea  coal  price,  the  anthracite 
prices  are  as  fixed  by  the  President  on  Aug.  23,  1917.  The  present  realiza- 
tion, all  companies  and  all  sizes,  including  washery  coal  and  both  the  labor 
increases,  is  calculated  to  average  $5.13  per  ton,  while  the  bulk  line  of  the 
chart  shown  in  Fig.  3,  plus  the  Nov.,  1918,  labor  increase,  would  be 
$5.32. 

The  capital  invested  per  ton  output  in  the  larger  and  better  equipped 
collieries  ranges  from  $5  to  $11,  with  an  average  investment  of  from  $7.50 
to  $8. 

PRICES  FIXED  BY  THE  PRESIDENT,  AUGUST  23,  1917 


WHIT: 

a  ASH 

RED 

ASH 

LYKENS 

VALLEY 

Com- 
pany 

Inde- 
pendent 

Com- 
pany 

Inde- 
pendent 

Com- 
pany 

Inde- 
pendent 

Broken  

$4.55 

$5.30 

$4.75 

$5.50 

$5.00 

$5.75 

Eee 

4  45 

5.20 

4.65 

5.40 

4.90 

5.65 

Stove 

4  70 

5  45 

4  90 

5.65 

5.30 

6.05 

Chestnut 

4  80 

5  55 

4  90 

5.65 

5.30 

6.05 

Pea              

4.00 

4.75 

4.10 

4.85 

4.35 

5.10 

20 


COAL  MINING  COSTS 
FIXED  PRICES,  DECEMBER  31,  1918 


WHITE  ASH 

RED  ASH 

LYKENS  VALLEY 

Com- 
pany 

Inde- 
pendent 

Com- 
pany 

Inde- 
pendent 

Com- 
pany 

Inde- 
pendent 

Broken  

$5.95 
5.85 
6.10 
6.20 
4.80 

$6.70 
6.60 
6.85 
6.95 
5.55 

$6.15 
6.05 
6.30 
6.30 
4.90 

$6.90 
6.80 
7.05 
7.05 
5.75 

$6.40 
6.30 
6.70 
6.70 
5.15 

$7.15 
7.05 
7.45 
7.45 
5.90 

Eee 

Stove       

Chestnut  
Pea  

The  prices  fixed  by  the  President,  Aug.  23,  1917,  are  given  in  the 
accompanying  table.  No  price  was  fixed  on  sizes  smaller  than  pea,  which 
was  decreased  60c.  per  ton  Oct.  1,  1917.  There  was  a  general  increase  of 
35c.  per  ton  Dec.  1,  1917,  and  one  of  $1.05  per  ton  Nov.  1,  1918.  Sizes 
smaller  than  pea  were  limited  to  a  maximum  50c.  per  ton  below  pea  coal 
by  order  of  Nov.  15,  1918. 

AVERAGE  COST  PER  TON,  DECEMBER,  1917,  TO  MAY,  1918 


Fresh- 
mined  coal, 
35,256,550 
Tons 

Washery 
Operations, 
3,431,916 
Tons 

Total,  In- 
cluding 
Washeries, 
38,688,466 
Tons 

Labor  

$2  593 

$0  687 

$2  423 

Supplies 

0  616 

0  260 

0  584 

Transportation,  mine  to  breaker  
Royalty,  current 

0.004 
0  153 

0.007 
0  102 

0.004 
0  148 

Royalty,  advance  

0.002 

0.002 

Depletion 

0  099 

0  077 

0  097 

Amortization  of  cost  of  leasehold  
Depreciation  .  . 

0.014 
0  091 

0.024 
0.86 

0.016 
0.090 

Pro  rata  suspended  cost  of  stripping  .... 
Contract  stripping  and  loading  

0.023 
0.009 

0.021 
0.009 

Taxes,  local 

0  054 

0  034 

0.052 

Insurance,  current  

0.016 

0.014 

0.016 

Insurance,  liability                             .    .    . 

0  058 

0.018 

0.055 

Officers'  salaries  and  expenses 

0  030 

0  019 

0  029 

Office  salaries  and  expenses  

0.048 

0.024 

0.045 

Legal  expenses 

0  005 

0.003 

0.005 

Miscellaneous  

0.026 

0.023 

0.026 

Total 

$3  841 

$1  378 

$3.622 

Increase  over  May  to  November,  1917  .. 

0.764 

0.365 

0.719 

MINING  COSTS 


21 


The  present  fixed  prices  Dee.  31,  1918,  per  ton  of  2240  Ibs.  f .o.b. 
mines,  are  given  in  the  accompanying  table.  Smaller  than  pea  is  not  to 
be  sold  within  50c.  of  maximum  pea-coal  price.  Thus  the  selling  price 
of  anthracite  has  been  increased  but  30.5  per  cent  over  the  pre-war  price, 
while  the  cost  of  production  has  gone  up  52  per  cent,  the  difference  having 
been  absorbed  by  the  operators. 

The  average  cost  as  reported  for  the  six  months,  Dec.,  1917,  to  May, 
1918,  inclusive,  prior  to  the  increase  of  Nov.  1,  1918,  but  including  that 
of  Dec.  1,  1917,  is  given  in  the  accompanying  table. 

Chief  among  the  factors  causing  fluctuations  in  mining  costs 
are  the  changes  in  wage  scales  and  the  tonnage  produced.  The 
labor  cost  per  ton  forms  70  to  80  per  cent  of  the  total  f.o.b. 
mine  cost.  In  73  mining  districts  for  which  detailed  statistics 
for  1918  were  published  by  the  Federal  Trade  Commission  the 
distribution  was  as  follows: 


• 

Per  Cent  of  f.o.b. 

Number 

1918 

Per  Cent  of 

Mine  Cost  that 

of 

Production, 

Total 

goes  to  Labor 

Districts 

Tons 

Tonnage 

60  to  64 

1 

247,000 

65  to  69 

10 

175,880,000 

35 

70  to  74 

33 

158,877,000 

32 

75  to  79 

18 

129,913,000 

26 

80  to  84 

11 

32,502,000 

7 

Totals 

73 

497  419  000 

100 

It  will  be  noted  that  the  difference  between  the  labor  cost 
proportions  from  district  to  district  are  much  less  than  between 
the  total  costs  themselves,  which,  excluding  lignite,  ranged  from 
$1.62  per  ton  in  a  West  Virginia  field  to  $4.45  in  an  Arkansas 
field. 

The  principal  reasons  for  the  difference  in  cost  of  the  various 
fields  are  the  physical  conditions  under  which  mining  must  be 
carried  on,  chief  of  which  is  the  thickness  of  seam  and  the  extent 
of  the  use  of  modern  machinery  for  mining  and  transporting 
coal  to  the  mouth  of  the  mine,  as  compared  with  the  old  fashioned 
pick  mining  and  mule  haulage.  It  is  a  mistake,  however,  to 
try  to  measure  the  advantage  one  district  has  over  another  by 
a  direct  comparison  of  labor  or  even  total  f.o.b.  mine  costs. 
Allowance  must  be  made  for  the  much  heavier  investment  neces- 


22  COAL  MINING  COSTS 

sary  In  the  fields  where  machinery  is  used  to  cut  down  the 
manual  labor  for  mining  and  transporting  the  coal. 

Nor  does  it  necessarily  follow  in  a  given  field  that  the  thicker 
the  seam,  the  lower  will  be  either  the  labor  cost  or  the  total 
f.o.b.  mine  cost  per  ton.  The  analyses  of  cost  by  thickness  of 
seam  mined  shown  in  the  Federal  Trade  Commission  reports 
indicate  that  after  a  certain  thickness  of  seam  is  reached — in 
most  districts  between  five  and  six  feet — the  mining  of  thicker 
seams  involves  higher  costs.  In  other  words  both  labor  and  total 
f.o.b.  mine  costs  per  ton  in  a  given  field  are  likely  to  decrease 
as  the  thickness  of  seam  increases  from  two  feet  up  to  between 
five  and  six  feet,  and  then  to  rise  as  the  thickness  increases 
still  further.  Apparently  in  such  cases  it  is  the  greater  amount 
of  labor  and  supplies  required  in  timbering  the  thicker  seams 
which  increases  the  costs. 

Wage  scales  for  each  field  are  fixed  with  relation  to  the  par- 
ticular mining  conditions  of  the  field.  Common  or  uniform 
increases  in  existing  wage  scales,  however,  have  widely  different 
results  in  their  effect  on  the  per  ton  labor  costs  of  the  different 
fields.  Thus  it  was  that  the  wage  increase  granted  in  November, 
1917,  for  which  a  uniform  price  increase  of  45c.  per  ton  was 
allowed,  increased  the  cost  about  28c.  per  ton  in  the  Illinois 
mines  of  F.  S.  Peabody,  according  to  his  testimony  before  Senator 
Reed's  committee.  On  the  other  hand,  it  has  been  found  that 
this  wage  increase  in  some  other  fields  increased  the  labor  cost 
as  much  as  70c.  per  ton. 

The  second  important  factor  in  causing  changes  in  the  per 
ton  cost  is  the  fluctuation  in  the  tonnage  produced.  Obviously 
the  greater  the  divisor,  the  less  per  ton  will  be  the  regular 
upkeep  expenses — whether  in  supplies  or  in  general  overhead 
charges.  But  also  the  proportion  of  so-called  "  non-productive  " 
labor  employed  in  the  mine  is  so  large  with  relation  to  the  labor 
paid  on  a  per  ton  basis  that  an  increase  in  the  production  will 
often  materially  decrease  the  total  labor  cost.  In  fact  the  increase 
in  production  may  be  so  great  as  to  obscure,  for  a  time,  the 
direct  effect  of  an  increased  wage  scale. 

There  is  little  authentic  information  available  as  to  costs  prior 
to  1916.  There  is  reason  to  believe  that  for  the  period  immediately 
preceding  the  war,  while  costs  had  been  gradually  increasing 
(disregarding  the  effect  of  fluctuations  in  production),  there  was 


MINING  COSTS 


23 


no  sudden  jump,  the  increase  taking  place  in  the  wages  from 
time  to  time  having  been  relatively  small  as  compared  to  total 
cost.  The  1916  costs,  therefore,  can  be  regarded  as  high  water 
mark  for  a  number  of  years  previous.  The  experience  of  the 
anthracite  field,  where  published  labor  costs  are  available  as 
far  back  as  1913,  supports  this  conclusion.  There  labor  costs 
on  fresh-mined  coal  of  operators  who  produced  about  60,000,000 
tons  annually  were  $1.62  per  gross  ton  in  1913,  $1.62  in  1914, 
$1.63  in  1915  and  $1.75  in  1916. 


3.00- 


FIG.   4. — Production   costs   in   Southwestern   Pennsylvania   for  the   years 

1916  to  1920. 

The  accompanying  charts,  Figs.  4  to  9,  inclusive,  for  all  of 
the  principal  producing  fields  in  the  United  States — these  fields 
produced  about  275,000,000  tons  in  1918 — show  the  rapid  rise 
of  costs  since  1916  and  also  give  some  measure  of  the  distribution 
of  costs.  The  figures  for  1916-1918  are  taken  from  the  Federal 
Trade  Commission  reports,  those  for  1919  and  1920  from  the 
reports  made  by  operators  to  the  National  Coal  Association,  and 
tabulated  by  the  Senate  Committee  on  Reconstruction  ("Calder 
committee").  The  allocation  of  these  costs  to  labor,  supplies 
and  general  expense  for  January -June,  1920,  has  been  compared 
on  the  basis  of  the  distribution  shown  in  the  Federal  Trade  Com- 
mission bulletins  which  covered  the  first  half  of  1920.  The 
fields  or  districts  are  those  established  by  the  Engineer  Committee 
of  the  Fuel  Administration,  and  are  defined  as  follows : 


24 


COAL  MINING  COSTS 


(1)  Southwest  Field,  Pennsylvania:  The  counties  of  Allegheny,  West- 
moreland, Fayette,  Greene  and  Washington,  in  the  State  of  Pennsylvania, 
except  (1)  that  portion  of  Allegheny  County  from  the  lower  end  of  Taren- 


8.00 


FIG.  5. — Production  costs  in  the  Indiana  district  for  the  years  1916  to  1920. 


8.00 


FIG.  6. — Production  costs  in  the  Illinois  No.  6  district  for  the  years  1916  to  1920 

turn  Borough  north  of  the  county  line;  (2)  the  territory  in  Westmoreland 
County  from  a  point  opposite  the  lower  end  of  Tarentum  Borough,  north 
along  the  Allegheny  Kiver  to  the  Kiskiminitas  Eiver,  along  the  Kiskiminitas 
Eiver  eastward  to  the  Conemaugh  Eiver,  and  continuing  along  the  Cone- 


MINING  COSTS 


25 


maugh  Kiver  to  the  county  line  of  Cambria  County;  (3)  operations  on 
Indian  Creek  in  Westmoreland  County;  and  (4)  operations  in  the  Ohio 
Pyle  district  of  Fayette  County.  See  Fig.  4. 


8.00 


FIG.  7. — Production  costs  in  the  Ohio  No.  8  district  for  the  years 
1916  to  1920. 


FIG.  8. — Production  costs  in  the  Pocahontas  Field  for  the  years 
1916  to  1920. 

(2)  Central   Field,   Pennsylvania:  The   counties   of    Tioga,    Lycoming, 
Clinton,  Center,  Huntingdon,  Bedford,  Cameron,  Elk,  Clearfield,  Cambria, 


26 


COAL  MINING  COSTS 


Blair,  Somerset,  Jefferson,  Indiana,  Clarion,  Armstrong,  Butler,  Mercer, 
Lawrence  and  Beaver,  and  operations  in  Allegheny  County  from  the  lower 
end  of  Tarentum  Borough  north  to  the  county  line,  and  in  Westmoreland 
County  from  a  point  opposite  the  lower  end  of  Tarentum  Borough  north 
along  the  Allegheny  Eiver  to  the  Kiskiminitas  Eiver  and  along  the  Kiski- 
minitas  Eiver  eastward  to  the  county  line  of  Cambria  County,  operations  on 
the  Baltimore  &  Ohio  E.E.  from  the  Somerset  County  line  to  and  including 
Indian  Creek  and  the  Indian  Creek  Valley  branch  of  the  Baltimore  &  Ohio 
E.E.  See  Fig.  9. 

(3)  Pocahontas  Field,  West  Virginia:  Operations  on  the  Norfolk  & 
Western  Ey.  and  branches  west  of  Graham,  Va.,  to  Welch,  W.  Va.,  including 
Newhall,  Berwind,  Canebrake,  Hartwell  and  Beech  Fork  branches;  also 
operations  on  the  Virginian  E.E.  and  branches,  west  of  Eock  to  Herndon, 
W.  Va.  See  Fig.  8. 


3.EO 


annggnnnnnnanoggnDnonnnnnanannnnnnnnnnnannnnnnDcinrannnnnnn 


FIG.  9. — Production  costs  in  the  Central  Pennsylvania  district  for 
the  years  1916  to  1920. 


(4)  District  No.  8,  Ohio:  The  County  of  Monroe,  the  County  of  Bel- 
mont,  except  the  township  of  Warren  and  operations  in  the  8-A  vein  in 
Flushing  and  Union  Townships,  the  County  of  Harrison  except  the  town- 
ships of  Monroe,  Franklin,  Washington  and  Freeport,  and  the  County  of 
Jefferson  except   the  townships   of  Brush  Creek,   Saline,   Eoss,   Knox  and 
Springfield.    See  Fig.  7. 

(5)  Bituminous   Field,  Indiana:  Coal  mined   in  the   State  of  Indiana 
other  than  Brazil  Block  coal.     See  Fig.  5. 

(6)  District    No.    6,    Illinois:  Including    Marion,    Jefferson,    Franklin, 
Williamson,  Johnson,  Hamilton,  Saline,  White,  Gallatin,  and  mines  along 
the  main  line  of  the  Illinois  Central  E.E.  between  Vandalia  and  Carbondale 
in  Clinton,  Washington,  Perry  and  Jackson  Counties.     See  Fig.  6. 


MINING  COSTS 


27 


The  charts  show  clearly  what  has  happened  to  costs  since 
1916  in  some  of  the  principal  fields  of  the  United  States.  The 
increases  in  these  fields,  based  on  the  1916  cost,  are  as  follows : 

COSTS  PER  NET  TON 
(Federal  Trade  Commission  Figures  Used  Exclusively) 


PENNSYLVANIA 

WEST  VIRGINIA 

Southwest 

Central 

Pocahontas 

New  River 

Labor 

F.o.b. 

Mine 

Labor 

F.o.b. 

'  Mine 

Labor 

F.o.b. 

Mine 

Labor 

F.o.b. 
Mine 

1916  (base)  
Jan.-March,   1920.  .  . 
April-  June,     1920.  .  . 

$0.82 
1.50 
1.88 

$1.19 
2.13 
2.62 

$0.92 
1.97 
2.17 

$1.32 
2.56 
2.82 

$0.56 
1.31 
1.51 

$0.87 
1.89 
2.11 

$0.74 
1.79 
1.85 

$1.00 
2.44 
2.39 

PER  CENT  OF  INCREASES  OVER  1916 


Jan.-March,  1920.  .  . 
April  -June,  1920.  .  . 

84 
130 

179 
120 

104 
136 

94 
114 

138 
175 

118 
143 

142 
150 

124 
120 

COSTS  PER  NET  TON 


OHIO 

INDIANA 

ILLINOIS 

No.  1 

No.  8 

Bituminous 

No.  3 

No.  6 

Labor 

F.o.b. 

Mine 

Labor 

F.o.b. 

Mine 

Labor 

F.o.b. 

Mine 

Labor 

F.o.b. 
Mine 

Labor 

F.o.b. 
Mine 

1916  (base)  .... 
Jan.-Mar.,  1920 
Apr.-June,  1920 

$0.84 
1.65 
1.64 

$1.17 
2.27 
2.09 

$0.78 
1.54 
1.79 

$1.02 

2.14 
2.47 

$0.87* 
1.63 
1.94 

$1.09 
2.00 
2.41 

$0.89t 
1.57 
1.77 

si.iot 

1.96 
2.21 

$0.86* 
1.63 
1.85 

$1.07* 
1.98 
2.29 

PER  CENT  OF  INCREASES  OVER  1916 


Jan.-Mar.,  1920 
Apr.-June,  1920 

97 
96 

94 

78 

97 
130 

110 
142 

87 
123 

83 
121 

77 
100 

78 
101 

89 
115 

85 
114 

*  April-December,  1916. 
t  July-December,  1916. 


In  the  foregoing  table  the  1920  figures,  while  not  strictly 
comparable  because  not  obtained  from  the  same  operators  as  the 
1916  figures,  are  probably  representative  enough  to  show  in  a 


28  COAL  MINING  COSTS 

general  way  the  change  in  conditions  since  1916.  If  the  Septem- 
ber, 1920,  total  cost  figures  of  the  Calder  committee  be  compared 
with  the  1916  total  f.o.b.  mine  cost,  the  increases  shown  would 
be  yet  more  marked,  as  the  effect  of  the  wage  increase  late  in 
the  summer  of  1920  was  to  increase  costs.  Such  increase,  how- 
ever, cannot  be  as  exactly  measured  because  the  1916  figures 
are  "revised"  costs  and  exclude  selling  expense,  while  the  1920 
figures  are  "reported"  costs,  and  include  selling  expense,  etc. 

As  the  events  of  the  past  few  years  have  shown,  labor  in 
coal  mines,  just  as  on  railroads,  holds  the  strategic  advantage 
of  being  able  to  tie  up  the  whole  country  through  an  effective 
strike,  it  is  not  likely  that  .costs  will  be  materially  lessened 
through  any  immediate  reduction  of  wages.  On  the  other  hand, 
the  necessary  writing  off  of  some  of  the  heavy  investment  charges 
caused  by  high  cost  development  during  the  past  few  years,  in 
order  to  get  the  investment  down  to  present  day  values,  will  in 
many  cases  increase  the  overhead  charges. 

Method  of  computing  tax  returns. — The  intent  of  the  law, 
according  to  an  article  on  this  subject  in  Black  Diamond,  is 
clearly  that  the  cost  of  the  coal  in  the  ground  shall  be  considered 
as  part  of  the  cost  of  the  same  coal  when  removed  and  sold.  The 
regulations  fail  to  carry  out  that  intent,  because  of  the  unwar- 
ranted assumption  that  all  the  tons  of  coal  in  the  mine  cost  the 
same  amount  per  ton.  This  is  what  gives  the  Treasury  Depart- 
ment's method  its  simplicity,  but  at  the  same  time  robs  it  of 
its  reasonableness,  because  the  assumption  is  contrary  to  fact.' 

It  is  self-evident  that  a  ton  of  coal  near  the  surface  or  exposed 
by  present  workings  is  worth  much  more  than  a  ton  buried  far 
down  in  the  earth  that  cannot  be  removed  for  many  years.  If 
no  improvements  in  mining  methods  were  expected,  the  value  of 
deeply  buried  coal  would  be  further  reduced  by  the  greater 
expense  that  would  be  required  to  bring  it  to  the  surface.  There 
is  a  possibility,  of  course,  that  improvements  in  methods  may 
keep  pace  with  the  difficulties  encountered  in  the  majority  of 
cases. 

But  whether  or  not  inventions  may  be  expected  to  offset  in 
some  measure  the  increasing  difficulties  of  greater  depth,  there 
is  nothing  to  compensate  for  the  time  element.  And  the  time 
element  is  always  an  important  factor.  No  extensive  deposit, 
whether  coal  or  ore  or  other  minerals,  can  be  mined  in  a  day 


MINING  COSTS  29 

or  a  year.  Almost  invariably  before  a  deal  in  mining  properties 
is  consummated  the  purchaser  has,  through  the  aid  of  specialists, 
made  exhaustive  studies  as  to  the  extent  and  quality  of  the 
deposit,  and  the  most  economical  methods  of  its  exploitation. 

It  would  not  be  economical  to  mine  one  ton  a  year,  and  it 
would  ordinarily  be  impossible  to  mine  the  whole  deposit  in  a 
year.  But  a  plan  is  adopted  between  these  extremes  usually 
based  on  the  annual  production  of  a  certain  definite  quantity, 
and  the  equipment  and  machinery  to  be  provided  in  order  to 
maintain  that  output  is  elaborated.  Ordinarily  the  purchaser 
not  only  has  these  plans,  but  has  concrete  evidence  of  them  that 
would  convince  any  fair-minded  and  disinterested  person  that 
in  purchasing  the  property  the  price  he  was  willing  to  pay  was 
based  upon  these  engineering  reports.  And  in  arriving  at  that 
price  he  always  does,  either  mathematically  or  intuitively,  take 
into  consideration  the  time  that  must  elapse  before  his  invest- 
ment can  be  realized  in  cash  by  operations. 

Let  us  suppose  that  the  mine  is  known  to  contain  1,000,000 
tons  of  coal,  and  that  the  plans  call  for  the  mining  of  20,000 
tons  per  annum,  indicating  a  life  of  50  yr.  for  the  mine,  and 
that  with  these  facts  in  mind  the  coal  lands  are  purchased  for 
$100,000.  This  is  the  cost  of  the  entire  deposit.  It  indicates 
the  average  cost  per  ton  is  10c.,  but  although  the  average  is  10c., 
that  figure  does  not  apply  to  the  tons  near  the  mouth  nor  to 
those  deeply  buried. 

To  determine  the  cost  of  the  several  tons  let  us  indicate  by 
V  the  value  of  a  ton  exposed  and  minable  to-day,  and  assume 
an  interest  rate  of  6  per  cent,  and,  to  avoid  unnecessary  intrica- 
cies of  computation,  let  us  further  assume  that  each  year's 
production  comes  at  the  end  of  the  year. 

The  cost  of  the  coal  to  be  mined  the  first  year  would  be 


The  cost  of  the  coal  to  be  mined  the  second  year  would  be 


The  cost  of  the  coal  to  be  mined  the  third  year  would  be 
20,0007(^5},  etc.  ; 


30  COAL  MINING  COSTS 

The  sum  of  this  series  for  50  terms  constitutes  the  cost  of  the 
entire  deposit.     Therefore: 

+ 


1.06  '  (1.06)2  '  (1.06)3  '  (1.06)' 

.  It  is  at  once  seen  that  the  series  in  the  bracket  is  equivalent 
to  the  present  value  of  an  annuity  of  $1  for  50-yr.  at  6  per 
cent,  which  is  readily  computed  at  $15.761.  Our  equation  is 
therefore  reduced  to: 

$100,000  =  20,0007(15.761). 

Solving  for  F,  we  find  that  the  cost  of  one  ton  minable  to-day 
is  31.72c. 

From  this  the  cost  of  the  20,000  tons  to  be  mined  each  year  may 
be  computed  as  follows: 


Cost 

Cost  of 

Tonnage 

per  Ton 

20,000  Tons 

1st    20,000  tons  $0 

.3172^  1.06      =$0.2992 

$5984 

2nd    20,000  tons 

.3172-i-(1.06)2  =  0.2823 

5646 

3rd    20,000  tons 

.3172-f-(1.06)3  =  0.2865 

5326 

4th    20,000  tons 

.3172  -Ml.  06)4  =  0.2512 

5024 

5th    20,000  tons 

.3172-K1.06)5  =  0.2370 

4740 

10th  20,000  tons 

.3172-h(1.08)10=  0.1772 

3544 

15th  20,000  rons 

.3172-(1.06)15=  0.1324 

2648 

20th  20,000  tons 

.3172-h(1.05)20=  0.09897 

1979 

30th  20,000  tons 

.3172-=-(1.06)30=  0.05528 

1106 

40th  20,000  tons 

.3172-H1.06)40=  0.03088 

618 

50th  20,000  tons 

.3172-H1.06)50=  0.01725 

345 

If  this  table  were  filled  in  complete  for  the  50  yr.  the  last 
column  would  total  up  to  $100,000,  which  is  the  cost  of  the 
entire  deposit. 

If  an  interest  rate  of  8  per  cent  were  adopted  the  cost  of  a 
ton  minable  to-day  would  be  40.87c.,  and  the  cost  of  the  respect- 
ive groups  of  20,000  tons  would  range  from  37.84  to  0.871c.  per 
ton. 

Other  properties  besides  mines  are  acquired  for  lump  sums, 
and  it  is  conceded  in  those  cases  that  the  purchaser  has  the  right 
and  the  duty  to  analyze  his  cost  and  set  up  in  separate  accounts 
a  fair  apportionment  of  it.  Thus  a  taxpayer  may  buy  for  a  lump 
sum  a  going  store  with  all  the  assets  and  liabilities  attached  to 
it.  He  is  expected  to  apportion  this  cost  and  set  up  separately 
in  his  books,  the  land,  buildings,  furniture,  merchandise,  accounts 


MINING  COSTS  31 

receivable,  good  will,  accounts  payable,  etc.  The  apportionment 
must  be  fair,  and  the  net  total  must  agree  with  the  aggregate 
cost. 

As  another  example  a  merchant  buys  a  shipment  of  miscel- 
laneous hides  for  a  lump  sum.  He  then  sorts  them  into  numerous 
grades.  The  best  hides  may  be  worth  many  times  as  much  as 
the  poorest  ones.  He  is  expected  to  apportion  the  cost  on  the 
basis  of  quality  and  value.  The  apportionment  must  be  fair  and 
the  total  of  the  costs  thus  allocated  to  the  several  grades  must 
agree  with  the  aggregate  cost.  If  the  hide  merchant  sold  all 
the  poorer  grades,  but  had  the  best  hides  on  inventory  at  the 
end  of  the  year,  and  priced  them  at  the  average  cost,  there  is 
little  doubt  but  that  the  Treasury  Department  would  compel 
him  to  use  a  higher  figure  and  would  assess  an  additional  tax. 

With  the  mining  company  the  situation  is  reversed.  In  the 
nature  of  its  operations  it  mines  and  sells  first  the  most  accessible 
coal — the  coal  that  really  cost  it  most — and  is  then  asked  to 
carry  the  less  accessible  coal  on  its  balance  sheet  at  the  average 
cost.  The  law  permits  a  reasonable  allowance  for  depletion,  and 
any  mining  company  that  makes  a  reasonable  and  fair  appor- 
tionment of  the  cost  of  its  mineral  deposits  should  be  accorded 
the  same  fair  treatment  that  is  accorded  to  the  merchant  of 
hides. 

Comparison  of  costs  and  distribution  of  revenue  of  German 
and  Middlewestern  mines. — A  study  of  mining  costs  and  the 
distribution  of  revenue  at  the  mines  of  the  Westphalian  Syndi- 
cate in  Germany  as  compared  with  the  mines  of  Illinois  and 
Indiana  throws  some  interesting  light  on  these  questions. 

The  mine  operators  of  the  Westphalian  district  in  Germany 
suffered  from  severe  competition  resulting  from  overproduction, 
and  various  efforts  were  made  to  find  relief  as  early  as  1850. 
Price  agreements,  which  were  forbidden  by  the  German  law, 
were  disregarded,  notwithstanding  the  heavy  penalties  imposed 
for  violations.  Finally  in  1885  the  Westphalian  Syndicate  was 
established,  and  continues  to  the  present  date.  It  is  a  selling 
organization  without  any  property  and  only  a  nominal  working 
capital. 

Its  affairs  are  administered  by  an  official  who  has  no  finan- 
cial interest  in  the  mines  and  acts  as  chairman  of  a  board  made 
up  of  one  representative  from  each  participating  company.  The 


32 


COAL  MINING  COSTS 


function  of  the  syndicate  is  to  sell  the  product  of  the  mines, 
coke  ovens  and  briquetting  plants  and  to  allot  to  each  company 
the  tonnage  which  it  should  produce. 

Twice  each  year  an  estimate  of  the  probable  requirements  is 
made,  and  a  tonnage  is  allotted  to  each  company  based  upon 
previous  production  after  allowance  has  been  made  for  the  ton- 


1075 
1050 
IOZ5 

JTAM 

la-l 

/ 

/ 

/ 

975 

ocn 

i 

i 
i 

925 

'QAfl 

i 

I 

875 

ftCA 

I 

*N 



^ 

i 

825 

ttAA 

• 

'     EA 
Ml 

RNINGS  ILLINOIS  PICK 

NERS  WOULD  HAVE  HAL 
D  RUNNING  TIME  EQUAL 
ttOF  WESTPHALIA  MM 

jf._ 

''' 

HA 

m 

» 

y 

775 
750 

/ 

\ 

/ 

/ 

\ 

/ 

/ 

\ 

' 

/ 

675 

650 

€25 

/ 

575 
550 
525 
500 

/ 

4VERA 
ILLII\ 

T  /W/V««  EARNINGS 
OIS  PICK  Ml  NEKS 

?/• 

! 

' 

/ 

\ 

/ 



^ 

/ 

/ 

\ 

1 

^ 

/ 

475 

X 

7 

N 

450 

"  1901  B02   1903  1904  1905  1906  1907  1908  1909  1910   1911    I9IZ  191, 

YEAR 

FIG.  10. — Wage  losses  to  miners  due  to  intermittent  work. 

nage  of  companies,  such  as  railroad,  furnace  and  other  such 
corporations,  which  consume  a  part  of  their  own  production. 

On  May  1  each  company  is  notified  how  much  coal  it  will 
be  called  upon  to  furnish  during  the  second  half  of  the  calendar 
year,  and  each  mine  can  make  its  arrangements  for  the  most 


MINING  COSTS 


33 


economical  production  of  the  tonnage  called  for.  Any  company 
falling  short  in  its  supply,  if  market  conditions  continue  as 
anticipated,  must  pay  damages  for  the  shortage  unless  the  deficit 
can  be  made  up  by  another  company. 


«v 
125 
2.00 
21.75 

ILK 

125 
1.00 
075 

f\Cf\ 

* 

^^ 

x' 

x^ 

^*^^ 

•»*.r 

X 

^* 

DiV/l 

r- 

^, 

?f/«J 

H 

—  — 
7 

«-•— 

,*^^ 

_^-* 

x^"' 

,-^- 

— 

——• 

^ 

^ 

... 

**" 

nATL 

"RIAL 

5,7> 

XES 

//vr/ 

1    i 

~R£STAND  6EI 

Vf>?> 

i£X 

YAtf 

-5/ 

s 

—  — 

_—  V 

^~ 

••I  -^ 

,,  •• 

—  -~ 

,•  "•* 

±^ 

**= 

^i^ 

•H^H 

-^ 

/ 

^\ 

nc 

I65i 

025 
0 

1694       1896       1896       1900        I90Z        1904       190^        1908 
YEAR. 

FIG.  11. — Distribution  of  gross  revenue  at  Westphalia  mines. 


1594     1596 


1900       190Z       1904      1906      1905 
YEAR 


1910       I9IZ 


FIG.  12. — Production  of  the  Westphalia  mines  compared  with  those 
of  Illinois  and  Indiana. 

Losses  due  to  inferior  preparations  are  borne  by  the  com- 
pany responsible  for  the  defect.  Prices  are  agreed  upon  and 
fixed  in  advance  semi-annually  and  take  into  account  the  quality 
of  coal  produced  from  each  mine,  making  it  immaterial  to  the 
purchaser  where  the  coal  comes  from,  because  of  the  adjustment 
of  price  to  the  intrinsic  value  of  the  material  sold. 

It  has  sometimes  happened  that  by  some  unforeseen  con- 
dition the  syndicate  was  not  able  to  market  through  its  ordinary 


34 


COAL  MINING  COSTS 


trade  channels  the  estimated  quantities  of  coal,  and  other  markets 
had  to  be  entered  in  order  to  permit  the  mines  to  operate  under 
the  most  economical  conditions.  Losses  due  to  these  difficulties 
are  borne  alike  by  all,  the  syndicate  paying  to  the  participants 
the  price  agreed  upon,  having  retained  a  commission,  from 
which  all  deficits  are  paid. 


JC.3 

son 

—  -  - 

ESTPHAUA 

^ 

'«. 

/ 

^s^ 

,/ 

>^ 

— 

275 

250 

225 
200 
175 
150 
I?R 

IL 

.INC 

IS 

2 

\ 

~~°***^. 

^-^**" 

L 

s 

•~-^> 

/ 

/ 

\ 
^ 

^—  • 

x 

*  —  , 

^-—  ' 

•         • 

'*«»• 

^ 

\ 

^ 

xv 

f 

/ 

\ 

x' 

\ 

/ 

"^ 

^ 

/ 

1 

1894     18%       1898        1900       1902 


1904 
YEAR 


1906       1908       1910       1913 


FIG.  13. — Comparison  of  the  average  number  of  shifts  worked  per  annum 
at  the  Westphalia  and  Middle  Western  mines. 


^.bU 

2.25 
2.00 
eU5 
5  1.50 
£  1.25 
1.00 
1.75 
,.50 

*v^^ 

STP 

Mil* 

/ 

f/ 

"'" 

^^ 

-— 

—  — 

***—" 

jS 

'^^ 

<^ 

INDIANA. 

L. 

S" 

*—  i 

*••** 

^^» 

Zi: 

'SC 

~^** 

+** 

^^ 

—IL 

L//VC 

394     1896       1898      1900       1902      1904      1906      1908      1910       191 

Y  EAR 

FIG.  14. — Average  price  per  ton  of  coal  at  Tipple,  Westphalia  and  Middle 

Western  mines. 

The  advantages  of  a  single  seller  marketing  50,000,000  tons 
of  coal  a  year  are  apparent.  Markets  are  available  to  the  syndi- 
cate which  individual  operators  could  not  reach.  Its  contracts 
are  made  for  five-year  periods,  and  this  assures  an  income  to  the 
operators  and  enables  them  to  finance  their  properties  and 
engage  in  business  which,  while  more  remunerative,  requires 
larger  investments.  Thus  they  have  erected  large  coking,  by- 
product and  briquetting  plants.  Such  financing  would  be  impos- 
sible with  the  uncertainties  of  ordinary  competition. 

The  higher  returns  have  made  possible  an  expenditure  of 


MINING  COSTS 


35 


money  for  improved  equipment,  safety  measures  and  labor-sav- 
ing devices  quite  unknown  in  this  country.  Complete  extraction 
of  coal  is  required  by  the  government,  and  it  is  estimated  that 
the  cost  of  flushing  to  sustain  overlying  strata  and  to  permit 
of  the  removal  of  all  coal  adds  25c.  per  ton  to  the  production 
cost. 


L-LO 

200 

C 

175 

Distribution  of 

i  Kfl 

B 

Revenue  1909 

B-lncfuctes  all  general 

f)  I7c 

salaries,  but-  notdepreci' 

I.C  7 

3 
) 

vl    flfl 

~c 

C 

-1 

~  ~~~ 

8itontirnerest  on  invest1*  ^ 
ment  or  sinking  fund, 

C-Represents  difference 

B 

B 

< 

between  gross  revenue    " 
and  operating  expense 

O*75 

~~ 

including  inte  reside- 
predariofysinh'ng  fvnd 

a 
i- 

and  nei  prof  its 
For  Jllinnis  2.6$ 

§ 

1 

V) 

UJ 

t, 

For  Indiana  0.8t 

g 

5 

« 

s 

? 

purpose. 

s 

1 

g 

^ 

* 

d 

S 

ft 

FIG.  15. — Comparison  of  revenue  at  Illinois,  Indiana  and  Westphalia  mines. 

The  coal  operators  are  enabled  to  provide  funds  for  the  pro- 
tection of  the  injured  employees  and  for  the  support  of  the 
families  of  those  fatally  injured.  They  also  provide  pensions  for 
the  incapacitated  and  the  aged.  The  cost  of  this  social  insurance 
in  1909  added  20c.  per  ton  to  the  cost  of  production. 

No  protest  has  been  made  by  the  consumer  against  the  higher 
coal  prices  which  have  followed  the  establishment  of  the  syndi- 
cate. The  increase  in  price  has  been  generally  accepted  as  the 
best  expedient  for  solving  a  most  vexatious  question.  Undoubt- 


36  COAL  MINING  COSTS 

edly  it  induced  more  care  and  economy  in  the  use  of  coal  and 
resulted  in  the  adoption  of  more  economical  engines  and 
improved  boiler  settings. 

The  Westphalia  production  increased  from  1,665,000  in  1850 
to  81,000,000  tons  in  1907;  at  the  same  time  the  number  of 
companies  was  reduced  from  100  to  76,  indicating  growth  of 
individual  companies  and  concentration  of  capital.  The  17  com- 
panies in  the  syndicate  the  output  of  which  was  sold  for  com- 
mercial use  and  which  were  not  allied  with  the  fuel-consuming 
industries  had  an  aggregate  annual  production  of  28,000,000  tons 
and  a  capitalization  of  $72,450,000  which  is  an  average  of  $4,200,- 
000  each. 

This  indicates  an  investment  for  plant  and  equipment  of 
$2.50  per  ton  of  annual  production.  The  capital  account  does 
not  include  any  outpay  for  coal  land,  as  all  the  coal  belongs  to 
the  government.  For  Illinois  the  capital  invested  in  1909  was 
$1.49  and  in  Indiana  $2.44  per  ton  of  annual  production.  This 
latter,  however,  includes  the  coal  rights,  which  represent  the 
major  portion  of  the  investment. 

The  accompanying  diagrams,  Figs.  10  to  15,  inclusive,  show 
graphically  the  points  developed  in  this  discussion. 

Conditions  where  operations  may  be  conducted  at  an  ap- 
parent loss. — Maintenance  charges  for  drainage,  timbering, 
ventilation  taxes,  etc.,  are  so  heavy  at  mines  that  it  may  be 
more  economical  to  continue  operations  in  the  face  of  an 
apparent  loss  than  to  shut  down  the  mines  entirely.  Particu- 
larly is  this  the  case  with  older  mines  having  long  underground 
hauls  and  high  pumping  heads.  Some  interesting  data  on  this 
subject  will  be  found  in  the  report  of  a  committee  appointed  to 
investigate  the  receivership  of  a  prominent  coal  corporation 
about  1911. 

This  committee  found  that  if  the  coal  properties  were  shut 
down,  the  annual  loss  will  be  $420,000.  If  they  were  operated  at 
the  standardized  cost  per  ton  of  $0.857  and  for  an  output  of 
3,000,000  tons  and  the  coal  sold  at  the  price  realized  the  previous 
year,  $0.8097,  the  loss  will  be  $141,900.  The  standard  cost 
includes  a  charge  for  interest  of  $0.067  and  for  depreciation  of 
$0.058,  a  total  of  $0.125  per  ton.  The  standard  costs  are  14.8 
per  cent  lower  than  1909-10  corresponding  costs,  17.4  per  cent 
lower  than  July  and  August,  1910,  corresponding  costs. 


MINING  COSTS  37 

This  short  report  was  amplified  into  the  following : 

The  coal  lands  have  been  injudiciously  acquired. 

Money  has  been  injudiciously  spent  in  equipping  the  plants. 
Overhead  charges  for  interest,  maintenance  and  depreciation 
are  therefore  high. 

The  current  market  selling  price  for  coal  was  so  low  as  to 
make  profitable  coal  mining  difficult,  if  not  impossible,  even 
if  the  coal  lands  had  been  secured  without  price,  and  had  been 
equipped  with  rigid  reference  to  economical  operation. 

To  shut  down  the  mines  and  wait  for  better  prices  would 
entail  an  annual  expense  for  power,  maintenance,  supervision, 
depreciation  and  interest  of  $420,000.  This  does  not  include 
an  annual  charge  of  $104,494  on  book  value  of  coal  lands  not 
immediately  identified  with  the  plants  to  be  operated. 

The  cost  of  mining  coal  if  operations  are  standardized,  will 
be  $0.857  per  ton  for  a  daily  output  of  12,000  tons,  a  monthly 
output  of  250,000  tons  and  a  yearly  output  of  3,000,000  tons. 

The  loss  from  continued  operation  will  depend  on  the  price 
obtained  for  coal  sold  as  follows: 


At  $0 . 66  loss  will  amount  to $561,000 

At  $0 . 70  loss  will  amount  to 420,000 

At  $0.70  loss  from  operations  and  loss  from  suspension  of 

operations  will  be  equal. 

At  $0. 79  loss  will  amount  to 200,000 

At  $0 . 8097,  price  netted  by  coal  sales  in  1909-10,  loss  from 

operation  will  be 141,900 

At  $0. 857  there  is  neither  loss  nor  profit  from  operation. 

At  $0.921,  profit  above  operation 192.000 

This  is  sufficient  to  pay  interest  on  obligation.     Coal  should 

therefore  continue  to  be  mined. 
At  $0.948,  profit  from  operation 272,0000 

This  pays  for  operation,  for  moneys  owed  and  for  present 
administration  charges, 

While  waiting  for  better  coal  prices,  costs  of  operations  were 
to  be  standardized  as  follows : 

By  revaluing  all  the  lands  and  equipment,  thus  reducing 
future  operating  overhead  charges. 

By  putting  the  management  in  the  hands  of  a  competent 
and  experienced  man  of  reliable  character. 


38  COAL  MINING  COSTS 

By  concentrating  operations  at  that  plant,  or  those  plants, 
where  coal  could  be  mined  most  cheaply. 

By  investigating  the  advantages,  if  any,  to  be  derived  from 
coking  the  product  of  these  mines. 

By  investigating  the  advantages,  if  any,  of  establishing  a 
washery  at  the  mines. 

In  making  its  investigations  the  committee  attempted  to 
determine  a  standard  cost  per  ton  of  mined  coal  for  a  standard 
output,  which  was  assumed  at  3,000,000  tons  each  year.  The 
standards  adopted  for  immediate  use  were,  per  ton : 

The  existing  wage  scale  for  mining  labor,  $0.485. 

Current  rates  of  wages  for  a  minimum  amount  of  other 
efficient  working  labor,  $0.175. 

Moneys  for  supervision,  supplies  and  other  bills,  taxes,  insur- 
ance, etc. ;  an  efficient  minimum,  $0.07. 

Depreciation  charges  based  on  revaluations,  on  experience, 
and  on  the  present  ascertained  coal  reserve  tributary  to  operat- 
ing plants,  $0.06. 

Interest  at  6  per  cent  per  annum  on  reappraised  values  of 
coal  reserves,  mining  buildings,  equipment,  etc.,  actually  used 
for  mining  operations,  $0.067. 

Other  expenses  not  standard  and  not  directly  appertaining 
to  mining  operations  were : 

Interest  and  other  charges  on  investments  at  present  inopera- 
tive, $0.029. 

Excessive  interest  load,  due  partly  to  investment  in  elaborate 
and  unnecessary  plants,  partly  to  deficits  accumulated  from 
former  years,  and  partly  to  other  causes,  $0.035. 

High  costs  of  administration  of  the  company's  business. 

COSTS  FOR  1910 

Operation $77,294  ' 

Maintenance 14,156 

General  expense,  excluding  insurance 37,912 

$129,362 
Less  allowance  for  mining  operation 48,000 


$81,362 
Cost  per  ton $0. 0271 

The  output  of  coal  can  fluctuate  from  no  tonnage,  if  the 
mines  are  closed,  to  a  maximum  daily  tonnage  of  17,000  tons. 


MINING  COSTS  39 

If  this  maximum,  of  17,000  tons  daily  could  be  attained  it 
would  reduce  mining  costs  about  as  follows: 

OUTPUT  PER  YEAR,  4,250,000  TONS 

Costs  per  Ton 

Mining  labor $0 . 455 

Other  labor 0. 15 

Operation 0 . 06 

Depreciation 0 . 06 

Interest 0 . 045 

Total $0.77 

TABLE  ON  BASIS  OF  3,000,000  TONS  ANNUALLY 
Daily  Output,  12,000  Tons 

Costs  per  Ton 

1.  Mining  labor $0.485 

2.  Other  labor 0. 175 

3.  Total  working  pay-roll  (1  and  2) $0.66 

4.  Operations $0. 07 

5.  Depreciation 0 . 06 

6.  Interest 0 . 067 

7.  Total  overhead  charge  (4,  5/6) $0. 197 


8.  Total  standard  cost  per  ton  of  coal  (3  and  7)  $0 . 857 

Systems  of  mining". — A  thorough  grasp  of  the  economics  of 
the  various  systems  of  mining  can  be  obtained  best  by  a  brief 
review  of  the  various  stages  of  development  that  have  lead  up 
to  the  adoption  of  the  systems  now  in  use.  A  study  of  the 
faults  that  were  found  in  these  older  methods  and  the  remedies 
that  were  applied  to  overcome  the  difficulties  is  of  prime  impor- 
tance. The  changes  in  the  systems  of  mining  in  the  Georges 
Creek  field,  one  of  the  oldest  bituminous  districts  in  the  country, 
was  described  in  a  paper  presented  before  the  West  Virginia 
Mining  Institute  in  1908,  from  which  the  following  has  been 
excerpted,  disregarding  the  methods  that  were  used  prior  to  1870 
when  there  was  apparently  little  attempt  at  systematized  effort. 

Fig.  16  illustrates  two  methods  followed  during  the  years 
between  1870  and  1880.  These  workings  are  inaccessible  to 
surveys  at  the  present  time  owing  to  the  creeps  and  squeezes 
induced  by  the  irregular  method  of  robbing  the  small  pillars. 
The  sketch  was  constructed  from  some  old  projection  drawings 


COAL  MINING  COSTS 


MINING  COSTS  41 

and  from  information  obtained  from  a  number  of  men  actually 
engaged  in  the  work.  The  main  headings  consisting  of  haulage 
road  and  airway  were  driven  on  the  strike  of  the  coal.  In  the 
first  method  the  room  headings  were  driven  in  pairs  from  the 
main  entry  at  intervals  of  600  ft.  and  on  the  rise  of  the  coal 
on  about  10-per  cent  grade.  From  these  headings  approximately 
25  rooms  were  driven  to  the  right  and  left  with  40-ft.  centers 
on  a  grade  of  4  per  cent,  giving  an  average  length  of  about 
350  ft.  The  rooms  were  14  ft.  wide  and  pillars  26  ft.  These 
pillars  were  found  to  be  totally  inadequate  and  extracting  them 
impossible.  Cross-cutting  the  pillars  at  frequent  intervals  was 
then  attempted  after  completion  of  the  rooms,  but  this  was 
generally  accompanied  by  creeps  closing  a  whole  district  at  a 
time.  The  maximum  height  of  the  superincumbent  strata  in  this 
territory  is  200  ft. 

The  second  method,  shown  in  Fig.  16,  was  adopted  later. 
The  maximum  thickness  of  the  overlying  strata  is  150  ft.  By 
this  method  headings  were  driven  from  the  main  entry  on  the 
rise  of  the  seam,  at  intervals  of  1000  ft.,  to  the  level  above,  and 
two  pairs  of  cross-headings  turned  to  the  right.  The  rooms 
were  driven  from  these  cross-headings  at  50-ft.  intervals  and 
14  ft.  wide,  leaving  a  pillar  of  36  ft.  The  length  of  rooms 
varied  from  300  ft.  to  550  ft.  These  pillars  were  also  of  insuffi- 
cient size;  robbing  was  conducted  spasmodically,  and,  although 
more  coal  was  recovered  than  in  the  adjoining  districts,  a  great 
deal  was  lost.  In  addition  to  the  small  pillars,  the  method  of 
robbing  them  was  calculated  to  promote  squeezes.  It  appears 
to  have  been  the  method  to  hold  the  strata  with  props  until 
sufficient  coal  had  been  removed  to  enable  the  weight  to  break 
the  props.  As  a  general  rule,  however,  before  this  was  attained 
the  weight  had  induced  a  creep,  which  is  well  known  to  have 
no  limits  within  a  territory  of  small  pillars. 

Fig.  17  represents  a  method  in  use  in  1890.  The  main 
entries  were  driven  from  the  slope  on  the  strike  of  the  seam, 
sufficient  grade  being  allowed  for  drainage.  Cross-headings  were 
driven  on  an  angle  of  about  35  degrees  to  the  main  entry  and 
headings  turned  off  these  parallel  to  the  main  entry.  Rooms 
were  turned,  as  shown,  from  all  headings  on  100-ft.  centers,  and 
pillars  split  by  half-rooms.  The  length  of  rooms  varied  from 
300  ft.  to  600  ft.,  and  were  15  ft.  wide,  leaving  pillars  42i/2  ft. 


42 


COAL  MINING  COSTS 


MINING  COSTS  43 

wide.  These  pillars  were  not  strong  enough  to  support  the 
overlying  strata,  500  ft.  high,  and  the  usual  creep  or  squeeze 
resulted  when  pillar  drawing  commenced. 

This  half  room  method  has  the  advantage  of  facilitating 
gathering  of  coal  and  doubling  the  support  of  the  haulway. 
The  squeeze  in  this  district  could  have  been  prevented  by  turn- 
ing the  rooms  from  the  haulway  on  200-ft.  centers  and,  after 
driving  the  half-rooms,  the  resultant  pillars  would  be  85  ft. 
wide.  While  this  would  have  avoided  a  squeeze,  the  great  weight 
to  be  supported  by  this  pillar  of  soft  coal  would  not  have  per- 
mitted a  very  high  percentage  of  recovery. 

Fig.  18  shows  a  method  adopted  in  1900.  The  maximum  dip 
is  15  per  cent,  and  the  greatest  thickness  of  superincumbent 
strata  425  ft.  The  slope,  together  with  parallel  air-course  and 
man-way,  are  sunk  on  the  heaviest  dip  of  the  coal,  and  double 
entries  turned  off  to  right  and  left  at  intervals  of  1000  ft. 
on  a  grade  of  1%  to  2*4  per  cent  in  favor  of  the  loads.  From 
these  haulways  cross-headings  are  deflected  at  intervals  of  240  ft. 
at  an  angle  of  about  25  deg.  and  driven  on  a  grade  of  4  per 
cent  to  7  per  cent.  Rooms  varying  in  length  from  100  to  800  ft. 
are  turned  on  the  rise  of  the  coal  from  these  cross-headings. 
The  rooms  are  driven  15  ft.  wide  on  65-ft.  centers,  leaving  pillars 
50  ft.  wide.  Twenty-five  rooms  are  driven  in  each  of  these 
diagonal  panels.  Unusually  large  protecting  pillars  are  left 
along  the  main  haulage  roads.  This  system  has  been  found  to 
be  especially  adapted  to  rapid  gathering  of  cars  thus  ensuring 
a  large  tonnage.  It  has  been  found,  however,  that  a  very 
large  recovery  from  the  pillars  is  impossible,  owing  to  the  many 
sharp  angles,  which,  in  a  thick  seam  of  soft  coal,  are  always 
difficult  and  ofttimes  impossible  to  extract.  This  sharp-angle 
method  was  even  resorted  to  formerly  in  cross-cutting  the  pillars 
preparatory  to  drawing  them,  but  this  has  been  changed  to  a 
rectangular  method,  thereby  increasing  the  actual  percentage 
of  pillar  coal  recovered  from  80  per  cent  to  83  per  cent.  The 
distance  of  rooms  apart  has  also  been  increased  in  the  last  few 
years  to  100-ft.  centers  giving  pillars  85  ft.  thick.  It  is  expected 
that  the  extraction  of  these  will  show  a  further  increase  in  the 
percentage  of  yield  from  pillars.  The  present  yield  from  head- 
ings, rooms,  and  pillars  under  this  system  is  about  90  per  cent, 


44 


COAL  MINING  COSTS 


MINING  COSTS  45 

considering  the  recovery  from  headings  and  rooms  at  100  per 
cent. 

Fig.  19  illustrates  a  method  instituted  in  the  latter  part  of 
1904.  The  main  haulway  is  an  extension  of  the  slope  from  the 
opposite  side  of  the  basin.  Double  entries  are  turned  off  from 
this  entry,  on  1%-per-cent  grade,  400  ft.  apart,  from  which 
rooms  are  driven  directly  on  the  rise  of  the  coal.  Rooms  are 
from  13  ft.  to  15  ft.  wide  and  practically  no  barrier  pillar 
left  between  the  room  face  and  the  air-course  of  the  panel  above. 
They  are  driven  at  100-ft.  intervals,  leaving  a  pillar  85  ft.  wide. 
The  length  of  a  panel  is  about  2500  ft.,  containing  22  rooms. 
There  are  five  such  panels  in  this  district  and  when  completed 
it  is  proposed  to  draw  the  pillars  in  a  retreating  fashion  with 
the  line  of  pillar  work  on  an  angle  of  45  deg.  across  the  whole 
district.  A  similar  method  in  another  district,  but  with  rooms 
on  a  deflection  of  35  deg.  from  a  right  angle,  is  yielding  SSy2 
per  cent  from  the  pillars  with  a  total  recovery  of  94  per  cent 
from  headings,  rooms,  and  pillars,  and  it  is  believed  that  this 
can  at  least  be  duplicated  if  not  exceeded  in  the  case  of  Fig  19. 
The  maximum  dip  in  this  district  is  6y2  per  cent  with  the 
greatest  height  of  the  overlying  strata  250  ft. 

With  this  resume  of  systems  used  at  different  periods  under 
conditions  now  known  in  view,  a  suggested  method  of  extracting 
the  coal  from  thick  soft  seams  with  a  brittle  top  and  a  height 
of  superincumbent  strata  of  400  ft.  or  less  is  presented  in  Fig. 
20.  The  general  design  of  the  mine  for  haulage,  drainage,  and 
ventilation  is  not  given,  because  they  are  variable  quantities, 
depending  on  conditions  which  change  with  the  locality,  and 
the  method  suggested  is  therefore  limited  to  the  ultimate  recov- 
ery of  the  coal.  By  this  method  a  territory  under  development 
is  divided  up  into  rectangular  panels  of  10  rooms  each.  The 
room  headings  are  driven  on  easy  grades  favorable  to  drainage 
and  haulage  and  the  panels  worked  in  pairs.  When  the  upper 
heading  has  been  driven  to  its  end  the  rooms  are  turned  at 
intervals  of  100  ft.  with  the  drawing  of  pillars  following  the 
retreating  method.  The  rooms  are  400  ft.  long  and  13  ft.  wide, 
leaving  pillars  87  ft.  in  width.  The  rooms  in  the  upper  panel 
are  limited  by  a  barrier  pillar  separating  them  from  the  head- 
ing above,  and  those  on  the  lower  panel  are  driven  through  to 
the  gob  of  the  upper  panel.  The  line  of  pillar  work  extending 


46 


COAL  MINING  COSTS 


MINING  COSTS 


47 


over  the  two  panels  should  have  an  angle  of  about  45  deg.  The 
length  of  rooms  can  be  varied  to  suit  the  conditions,  and,  when 
the  height  of  the  overlying  measures  exceeds  400  ft.,  the  thick- 
ness of  pillars  should  be  increased  accordingly. 


FIG.  20. — Suggested  method  of  working  thick,  soft  seams,  having  a 
brittle  top  and  400  ft.  cover. 

Fig.  21  shows  the  method  of  drawing  the  pillars  in  detail. 
The  rectangular  method  should  be  carefully  adhered  to  and 
all  sharp  angles  avoided.  When  the  work  commences  a  cut 
is  made  separating  a  block  of  coal  30  ft.  wide  and  87  ft.  long. 
This  piece  is  further  subdivided  into  blocks  varying  from 
6  X  12  ft- to  15  X  12  ft- in  size>  which  are  cut  off  and  extracted 
successively  as  shown.  In  no  case  should  the  small  blocks  cut 
off  exceed  in  size  the  distance  a  man  can  shovel  under  average 
conditions,  which  is  about  15  ft.  The  largest  of  the  small  blocks 


48 


COAL  MINING  COSTS 


should  be  removed  last  because  it  is  there  that  the  greatest  pres- 
sure manifests  itself  in  the  removal  of  the  original  block  cut 
from  the  pillar.  When  this  block  has  been  removed  another 
cut  is  made  and  the  process  repeated.  The  line  of  gob  should 
be  approximately  on  an  angle  of  45  deg.  If  it  is  found  impos- 
sible to  take  out  the  small  blocks  clean,  on  account  of  the  gob 
from  the  preceding  work  running  in  on  the  new  cut,  a  row  of 
props  can  be  set  on  the  upper  side  of  the  large  block  to  hold 
the  gob  when  the  small  blocks  below  are  removed.  This  method 


L. 


FIG.  21. — Method  of  drawing  pillars  to  be  used  in  the  system  shown  in 

Fig.  20. 

of  working  should  result  in  a  total  average  recovery  of  fully 
95  per  cent  of  the  seam  from  headings,  rooms,  and  pillars  and 
ensure  the  workings  against  creeps  and  squeezes. 

One  of  the  most  interesting  studies  of  costs  as  applied  to 
the  different  systems  of  mining  flat  seams  of  coal  was  included 
in  a  paper  presented  before  the  February,  1915  meeting  of  the 
Am.  Inst.  of  Mining  and  Metallurgical  Engineers,  excerpts  from 
which  are  given  herewith. 

It  may  be  stated  that  the  unit  with  which  we  have  to  deal  in 
studying  mining  costs  is  the  room ;  what  takes  place  at  its  face 
is  the  real  productive  work  of  the  mine,  and  all  else  under- 


MINING  COSTS 


49 


ground  is  for  the  purpose  of  serving  best  the  worker  at  the  room 
face.  Fig.  22  shows  several  typical  methods  of  procedure.  They 
are  of  particular  interest  in  that  one  may  see  them  in  mines 
following  the  same  plan,  working  the  same  seam,  under  con- 
ditions which  admit  of  comparison.  The  features  of  these 
methods  are  given  in  Table  I. 

In  all  of  these  methods  variations  may  be  seen,  from  entries 
driving  with  no  rooms  turned  to  entries  driving  with  two  or 
more  rooms  turned  and  driving  as  the  entries  advance ;  in  respect 
to  the  robbing,  one  may  see  variations  from  robbing  following 


•--Pillar-. 


-H  2/K-4ET-H ; 

I 


D  E  F 

FIG.  22. — Typical  methods  of  procedure  in  working  rooms. 


immediately  upon  the  completion  of  the  first  two  rooms  to  the 
robbing  following  at  an  indefinite  date  after  the  completion  of 
the  first  workings  of  the  panel.  Where  continuous  paneling,  or 
advancing  robbing,  is  in  effect,  robbing  is  not  compelled  to  wait 
until  the  completion  of  all  the  entries  of  the  panel. 

The  number  of  rooms  per  entry  varies  from  about  12  to  an 
indefinite  number,  and  the  depth  of  the  room  varies  from  about 
300  to  about  800  ft.  The  amount  of  timber  and  the  manner 
and  time  of  placing  same  depend  largely  upon  the  individual 
miner,  and  as  a  rule  there  are  no  specific  instructions  for  his 
guidance ;  also,  in  general,  no  effort  is  made  to  recover  the  timber 
in  robbing. 


50 


COAL  MINING  COSTS 

TABLE  I 

METHODS  OF  PROCEDURE  IN  DRIVING  ROOM 


Sketches 

A 

B 

C 

D 

E 

F 

Width  of  room  in 

feet  

24 

20 

20 

30 

36 

36 

Width  of  pillar  in 
feet  

36 

65 

40 

45 

54 

54 

Location  of  track  .  . 

In  center 
of  room 

Along 
robbing 
rib 

Along 
robbing 
rib 

Along 
robbing 
rib 

Along 
robbing 
rib 

Along 
robbing 
rib 

Location  of  gob  .  .  . 

Along 
both  ribs 

Opposite 
robbing 
rib 

Opposite 
robbing 
rib 

Between 
tracks 

Opposite 
robbing 
rib 

Between 
tracks 

Number  of  men  per 
room  

Ito2 
rooms 

1 

lor  2 

2 

6 

4 

Feet  of  room  face 

per  man  
Feet   of   entry  per 
man  

48 
120 

20 

85 

20  or  10 
60  or  30 

15 
37.5 

8.5 
15 

9 
22.5 

A  method  of  procedure  observed  at  one  mine  (but  which 
has  not  as  yet  been  sufficiently  tested  out  in  the  matter  of  recov- 
ering the  pillar  to  warrant  its  unreserved  adoption),  is  shown 
in  Fig.  22,  E.  Here  it  is  intended  that  rooms  shall  be  driven 
36  ft.  wide  on  centers  90  ft.  apart,  carrying  a  room  face  at  an 
angle  of  45  deg.  and  a  single  track  along  the  robbing  rib  but 
curved  to  parallel  and  follow  the  length  of  the  room  face.  It 
is  intended  to  work  six  men  to  the  room,  the  gathering  motor 
receiving  and  placing  three  cars  at  a  time.  Immediately  upon 
the  completion  of  the  room  the  pillar  is  to  be  withdrawn.  By 
this  method  of  procedure  a,  high  degree  of  concentration  will  be 
effected  and  the  efficiency  of  the  gathering  motors,  mining 
machines,  and  miners  will  be  increased.  It  is  also  hoped  that 
by  carrying  the  working  face  on  a  diagonal,  fewer  unexpected 
falls  of  top  will  occur  than  at  present,  because  the  fracture  will 
generally  be  partly  exposed  before  the  entire  coal  support  is 
removed  from  beneath  it. 

In  most  mines  the  miner  at  the  face  is  responsible  for  the 
safe  working  conditions  of  his  room.  In  the  above  methods  of 
procedure,  therefore,  one  might  say  that,  for  the  same  expendi- 
ture of  time,  energy,  and  watchfulness,  the  relative  degree  of 


MINING  COSTS  51 

security  which  the  miner  may  feel  as  a  result  of  his  efforts  is 
inversely  proportional  to  the  room  space  he  occupies.  It  is  also 
true  that  for  cars  of  the  same  height  the  energy  expended  by 
the  miner,  or  the  work  done  in  loading  the  coal,  is  much  less 
where  two  or  more  men  work  per  room  and  the  room  space  per 
miner  is  low,  than  where  one  man  works  per  room  and  the 
room  space  per  miner  is  high. 

In  the  grouping  of  rooms  as  outlined  in  Fig.  22  many 
arrangements  were  made  from  which  have  matured  certain  well- 
defined  plans.  Probably  the  consensus  of  opinion  favors  the 
panel  system,  but  even  with  it  there  are  differences  of  opinion. 
Many  men  think  that  the  entries  should  be  driven  to  the  inside 
lines  of  the  property  and  the  coal  extracted  retreating;  others 
think  that  half  of  the  property  should  be  worked  advancing 
and  the  remainder  retreating ;  yet  others  think  that  all  or  nearly 
all  of  the  coal  should  be  extracted  as  the  entries  advance.  It 
is  probable  that  it  is  best  to  extract  the  coal  in  such  a  manner 
that  the  present  worth  on  the  returns  from  the  mining  venture 
will  be  greatest,  both  to  the  property  owner  and  the  operator, 
leaving  only  such  coal  during  the  advance  of  the  entries  as 
will  permit  of  profitable  mining  until  the  final  exhaustion 
of  the  property.  Figs.  23  and  24  show  typical  plans  on  the 
panel  system.  Fig.  23  is  the  square  or  rectangular  panel,  Fig. 
24  the  continuous  panel. 

For  purposes  of  discussion  and  demonstration  a  typical 
property  of  1000  acres  of  the  approximate  shape  indicated  in 
Fig.  25  will  be  assumed  from  which  it  is  desired  to  produce 
2800  tons  per  day  when  running  at  maximum.  The  dip  is  2.5 
per  cent  and  the  strike  lies  at  an  angle  of  about  10  deg.  to  the 
long  axis  of  the  property,  its  general  direction  being  from  the 
upper  right  hand  corner  of  the  property  to  the  lower  left  hand 
corner.  The  coal  is  fairly  clean,  6  ft.  thick,  and  the  condition 
of  grades,  top  and  bottom,  are  fair.  It  is  also  assumed  that  the 
rate  of  loading  per  man  per  day  will  be  16  tons.  The  questions 
to  be  decided  are  what  method  of  procedure  and  what  plan  are 
best,  to  determine  which  the  following  information  is  desired : 

1.  What  period  of  time  will  be  required  to  reach  the  desired 
output  ? 

2.  How  many   day  laborers,   mining  machines,   mine   cars, 
mules  for  gathering,  and  main-haulage  motors  will  be  required? 


52 


COAL  MINING  COSTS 


t. 


Barrier  Pillar 


Barrier  Pillar 


$- 


-I 


ain  Errfry 


FIG.  23. — Typical  plan  of  mining  on  the  square  or  rectangular  panel  system. 


...A 


i   I  I   l  l  VI   I  I  I  I  I  \  I   I    IJ_U_J 


.  24. — Typical  plan  of  mining  on  the  continuous  panel  system. 


MINING  COSTS  53 

3.  How  much  main  entry,  main  entry  track  and  trolley  wire ; 
cross  main  entry,  cross  main  entry  track  and  trolley  wire ;  room 
entry,  room  entry  track,  and  rooms  and  room  track  will  be 
required  ? 

4.  What  is  the  length  of  the  average  car  haul? 

5.  What  is  the  relative  amount  of  power  for  ventilation? 

6.  What  is  the  acreage  of  standing  pillars,  the  estimated 
relative  cost  of  production,  and  the  estimated  percentage  of 
recovery  ? 

The  methods  referred  to  in  Fig.  22,  0,  and  the  plans  of 
mining  in  Figs.  23  and  24  will  be  applied  to  the  problem  as 
follows : 

First  Form. — Drive  the  third  entry  of  the  panel,  turn  the 
last  two  rooms  on  this  entry  first,  start  removing  the  pillar 
immediately  upon  the  completion  of  the  next  to  the  last  room, 
and  continue  to  drive  all  the  rooms  in  the  panel  only  fast  enough 
to  provide  for  the  uninterrupted  advance  of  the  robbing.  Work 
two  men  to  the  room  and  in  the  air-courses  and  on  the  pillars. 
Only  this  method  of  procedure  will  be  applied  to  the  square  and 
continuous  panels,  and  the  following  methods  to  the  square 
panel. 

Second  Form. — Drive  the  rooms  of  the  panel  as  they  are 
encountered,  turning  the  first  entry  of  the  panel  when  it  is 
reached,  and  start  robbing  immediately  upon  the  completion  of 
the  last  room  on  the  third  entry  of  the  panel.  Work  one  man 
to  each  working  place. 

Third  Form. — Drive  the  rooms  and  entries  of  the  panel  as 
they  are  encountered,  start  robbing  immediately  upon  the  com- 
pletion of  the  last  room  in  the  third  entry,  and  continue  the 
robbing  until  the  completion  of  the  panel.  Work  one  man  to 
every  other  room,  but  advance  all  rooms  and  work  one  man  to 
each  pillar. 

The  accompanying  table  shows  the  comparison,  as  well  as 
some  other  figures  to  which  further  reference  will  be  made. 

From  these  data  it  may  be  concluded  that  the  first  form  of 
procedure  and  the  plan  of  mining,  Fig.  24,  are  best. 

The  period  of  time  required  to  reach  the  desired  output 
was  determined  for  the  several  methods,  as  shown  in  Figs.  25, 
26,  27,  28  and  29;  the  location  of  the  working  faces  from  day 
to  day,  as  determined  by  the  assumed  rate  of  advance  of  16 


54 


COAL  MINING  COSTS 


tons  per  man,  was  plotted  on  a  map,  and  the  total  number  of 
faces  at  the  time  the  desired  output  was  reached  were  counted, 
from  which  data  the  tonnage  curves  were  plotted  as  shown  in 
Figs.  25A,  26A,  27A,  28A  and  29A. 


First 
Form 

Continuous 
Panel 

Second 
Form 

Third 
Form 

Ad- 
vancing 
Method 

Output  reached   months 

53 

42 

62 

92 

7 

Day  laborers 

60 

65 

82 

102 

31 

ADVANCING  METHOD  USES  8  ASSISTANT  FOREMEN 


Mining  machines          .                 

9 

9 

14 

18 

8 

Mine  cars 

275 

310 

335 

465 

155 

Mules 

18 

22 

24 

32 

10 

Motors 

4 

4 

5 

6 

2 

7  850 

6500 

9  300 

13  950 

600 

Main-entry  trolley  

Cross  main  entry  

Cross  main-entry  track  
Cross  main-entry  trolley  
Room  entry  and  room-entry  track  
Room  track                    .            

5,550 

10,700 
15.84C 

5,150 

12,700 
20,500 

8,850 

33,900 
96,800 

13,500 

50,400 
230,300 

1,000 

7,000 
18  100 

6,180 

5,333 

7  420 

10  230 

3  640 

Ventilation  power,  kilowatt  hours  

40 
62  8 

42 
65  7 

125 
168  2 

175 

277  0 

20 

13  8 

Relative  cost  of  production  

1.33 

1.24 

1.76 

2   1 

1 

Percentage  of  recovery   

94 

95.5 

83 

80 

97 

In  arriving  at  the  relative  number  of  men  and  mules 
required,  rates  for  performing  certain  tasks,  taken  from  time 
study  observations  were  used.  The  amount  of  rolling  stock 
required  is  based  on  the  assumption  that  the  equipment  will 
travel  at  the  same  rate  of  mileage  per  day ;  the  other  items  com- 
pared were  taken  direct  from  the  maps.  In  the  absence  of 
facts  for  comparison,  opinions  of  creditable  authorities  were 
sought  in  very  instance. 

These  methods  of  procedure  and  both  plans  of  mining  have 
been  designed  to  meet  certain  wants.  In  some  instances  certain 
features  of  the  plan  have  been  prescribed  by  the  land  owners 
in  order  to  safeguard  their  interests  from  ' '  squeezes ' '  and  losses 
of  coal  due  to  lack  of  proper  supervision.  Were  the  proper 
supervision  supplied  and  better  methods  of  procedure  adopted, 
the  restrictions  in  the  plan  of  mining  might  very  properly  be 


MINING  COSTS 


55 


COAL  MINING  COSTS 


removed.  Other  details  of  design  have  been  the  result  of  accept- 
ing certain  " rules  of  thumb"  which  have  since  been  proved 
wrong,  and  yet  other  details,  although  admittedly  wrong  and 
expensive,  have  been  introduced  rather  than  combat  the  wrongs 
which  they  are  designed  to  circumvent. 

In  the  plan  of  mining  shown  in  Fig.  23,  the  frequent  inter- 
position of  barrier  pillars  is  for  the  purpose  of  confining  a 
squeeze  and  limiting  its  range  of  destructive  action.  The  use 
of  these  barriers  is  imperative  under  the  methods  of  procedure 


3000 


2750 


2250 

^•2000 

o 
o 

,.1750 
& 


750 

500 
250 


Tofal  Tonnage 


"•'from  all  Places 


(Lfooms  and  ftobbingl"-? Flat 


I 


<-Rooms  and  fobbing  3& 'Flat 


Mam  Entry  and  First  F/crf. 


'^and  3^  Flat  Entries 
20       25        30 


60       65 


Months  of  25  Days  per  Month 

FIG.  25A. — Tonnage  curves  under  the  method  of  procedure  shown  in  Fig.  25. 

that  involve  large  areas  of  long-standing  pillars  and  where  the 
degree  of  supervision  is  low.  It  is  to  be  regretted  that  their 
use  is  so  common,  for  they  tend  to  interfere  seriously  with 
the  maximum  degree  of  concentration  because  one  is  seldom, 
if  ever,  able  to  provide  a  satisfactory  output  from  a  single  panel, 
and  then  only  for  a  short  period  of  time.  Where  two  or  more 
panels  are  required  to  produce  the  output,  the  further  the  work- 
ings advance  the  more  distantly  seperated  they  become,  or  other 
important  considerations  must  be  sacrificed.  Disadvantages  of 
the  unit-panel  plan  may  be  seen  in  the  curve  in  Fig.  30  which 
shows  the  great  variation  in  the  tonnage  obtained  daily,  vary- 


MINING  COSTS 


57 


58 


COAL  MINING  COSTS 


ing  from  zero  at  the  opening  of  the  panel,  augmented  by  a 
more  or  less  constant  rate  of  increase,  to  a  certain  maximum 
number  of  tons,  and  then  a  gradual  decline  to  zero  again.  If 
a  certain  number  of  tons  per  day  gathered  from  the  panel  is 
accepted  as  100  per  cent  efficiency  for  a  gathering  motor,  as 
shown  in  Fig.  30,  it  will  be  noticed  that  the  motor  is  at  first 
working  at  a  very  low  efficiency,  which  gradually  increases 
until  the  maximum  is  reached,  at  which  time  another  motor 
must  be  added,  and  the  average  efficiency  of  the  two  motors  ig 


^-, 

t* 

2250 

1500 
1250 
1000 
750 
500 
250 
0 

A 

f 

& 

cf 

/ 

4** 

A* 

/ 

/ 

f. 

x> 

U 

fc 

\6 

!n\  / 

S\ 

^ 

u 

M 

/ 

i 

A 

I 

\       , 

A 

A 

/ 

\ 

A 

\ 

\ 

1 

/f* 

!ah+ 

/ 

i 

I 

Y 

\ 

/ 

Y 

V 

\ 

\ 

29 

// 

\ 

2 

/ 

\ 

/\ 

/ 

\«, 

binew 

tflhjn 

te?w/ 

A 

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f-f^RiahH 

3 

_1 

A| 

A 

/ 

\  1    ^ 

\ 

5         10       15       20 

>5 
'   M< 

0 

>nths 

5       40        45       50       55       60       < 
at  25  Days  per  Month 

»5       70       75       f 

'O      85      9< 

FIG.  26A. — Tonnage  curves  under  the  method  of  procedure  shown  in  Fig.  26. 

about  50  per  cent;  there  is  a  similar  drop  in  efficiency  with 
each  motor  that  is  added,  until  the  maximum  tonnage  from  the 
panel  is  reached,  after  which  the  process  of  removing  motors 
from  the  panel  is  begun.  In  some  measure  this  degree  of  effi- 
ciency may  be  increased  by  working  the  motors  over  more  than 
one  panel,  as  is  often  done,  and  a  better  efficiency  curve  might 
be  obtained  more  nearly  in  accordance  with  the  full  line  shown, 
but  in  practice  a  rigid  watch  must  be  kept  on  this  detail,  or 
more  often  than  otherwise  a  lower  degree  of  efficiency  than  that 
shown  will  result. 

If,  in  the  preceding  paragraph,  instead  of  considering  the 


MINING  COSTS 


60 


COAL  MINING  COSTS 


efficiency  of  the  gathering  motor  the  efficiency  of  day  laborers 
or  the  tons  produced  per  unit  of  material  and  equipment  in 
use  had  been  considered,  the  same  general  discussion  would 
apply.  Where  low  efficiencies  are  obtained  from  day  laborers, 
material,  and  equipment,  low  efficiencies  are  also  obtained  from 
the  miners  at  the  room  face.  For  these  reasons  it  is  difficult, 
and  in  practice  well-nigh  impossible,  to  establish  any  constant 
relation  between  a  given  tonnage  desired  to  be  uniformly  pro- 


3000 


25      3O       35      4O      45        50      55       6O      65 
Months  oi  25  DCILJS  per  Month 

FIG.  27A. — Tonnage  curves  under  method  of  procedure  shown  in  Fig.  27. 

duced,  and  the  amount  of  material,  equipment,  and  day  laborers 
required  to  produce  that  tonnage;  the  efficiency  of  these  quan- 
tities rises  and  falls  with  the  rise  and  fall  of  the  tonnage  curve, 
although  in  an  erratic  manner. 

Thus  it  would  appear  that  the  square  panel,  while  designed 
to  meet  certain  requirements,  does  so  at  the  loss  of  much  that 
is  to  be  desired,  and  introduces  new  complications.  The  barrier 
pillars  are,  as  the  term  implies,  for  the  purpose  of  barricading 
against  some  impending  danger,  such  as  an  unforeseen  squeeze. 
Since  no  one  can  predetermine  where  or  when  these  squeezes 


MINING  COSTS 


61 


62 


COAL  MINING  COSTS 


will  occur  it  sometimes  happens  that  the  barrier  pillars  are 
provided  where  they  are  not  needed;  yet  experience  has  shown 
the  wisdom  and  necessity  of  their  use  under  certain  conditions. 


& 


f* 


-V 


V 


v 


i 


V 


\i 


Si 


• 


1    1    1    §    1 

he>Q  js 


1    i 


They  would  be  used  less  frequently  if  the  square  panels  were 
made  rectangular,  but  the  same  degree  of  security  would  not  be 
obtained  unless  the  entries  were  driven  to  the  limit  of  the 
rectangle,  with  few  or  no  rooms  driven  as  the  entries  advance. 
If  we  accept  it  as  axiomatic  that  when  a  room  is  driven  to  com- 


MINING  COSTS 


63 


64 


COAL  MINING  COSTS 


pletion  its  pillars  should  be  immediately  removed  in  order  to 
obtain  the  best  results,  or  that  it  is  equally  as  fundamental 
to  open  up  no  new  entries  until  ready  to  mine  from  them,  and 


13 


,1. 

=8 


\ 


§     I     §     §     2     §     » 

(n?q  jdcl  suoj. 


that  mining  should  then  be  conducted  at  the  maximum  rate  of 
production,  the  rectangular  panel  that  involves  either  long* 
standing  pillars  or  long-unproductive  entries  must  be  rejected. 

The  continuous  panel  obviates  the  necessity  for  frequently 
interposing  a  barrier  pillar  and  it  is  especially  well  adapted 


MINING  COSTS 


65 


to  a  property  where  the  main  entries  are  driven  to  the  dip. 
However,  the  tonnage  from  a  single  continuous  panel  is  limited, 
and  where  the  main  entries  of  a  property  go  to  the  rise  the 
maximum  degree  of  concentration  cannot  be  obtained  or  the 
rooms  off  the  cross  entry  will  go  to  the  dip.  Advancing  robbing 
is  impracticable  because  the  pockets  in  the  pillars  go  to  the  dip. 
The  rate  of  production  from  a  single  room  entry  rises  and  falls 
in  the  same  manner  as  the  rate  of  production  in  the  room  entries 
of  the  square  panel  and  the  general  discussion  above  in  reference 
to  the  square  panel  applies  to  the  continuous  panel. 


l*t>O 

1400 
1350 
1300 
1250 
1200 
1150 
1100 
1050 
1000 
950 
900 
650 

<§750 
u700 
g.650 
°-600 

00  1 

8n  " 

K. 

.\ 

/! 

\ 

f 

j* 

^ 

. 

1 

\ 

\ 

[J 

1 

1 

\ 

1 

V 

• 

1 

\ 

1 

\ 

\ 

/ 

V 

J500 

f>J 

\ 

. 

i 

400 
350 
300 

\ 

J 

s 

J 

\ 

250 

... 

% 

- 

•3 

S 

< 

^*- 

-*. 

•— 

^ 

N- 

^ 

J 

- 

\ 

150 
100 
50 

s* 

\ 

j 

» 

g 

k> 

v- 

y 

fT 

*-•  '.fficienci/  Curves 

\ 

40  % 

-fi 

S 

of- 

Ocrtherinq  Motors 

V 

C 

f^ 

\ 

1 

i 

1    2  2 

4   5  6  7   8  9  JO      12       14      16      18      20     22      24      26      28     30      3 

1     34 

Months 

FIG.  30. — Variation  in  the  tonnage  daily  obtainable  from  the  unit  panel 
and  the  efficiency  of  gathering  motors  working  in  the  panel  when 
proceeding  as  outlined  in  Fig,  26. 

However,  if  one  follows  the  history  of  the  development  of 
mining  methods  from  the  early-day  single-entry  system  to  the 
present-day  panel  system,  it  will  be  found  that  the  square  or 
nearly  square  panel  meets  sound  mining  practice  more  closely 
than  any  of  the  plans  which  have  preceded  it.  Until  methods 
of  procedure  are  adopted  which  make  the  restrictions  of  the 
panel  unnecessary,  or  until  a  plan  of  mining  is  devised  with- 
out the  objectional  features  of  the  panel,  but  retaining  its  may 
favorable  features,  the  square  panel  will  be  accepted  by  many 
operators  as  the  standard  plan  of  mining. 


66  COAL  MINING  COSTS 

For  many  years  it  has  been  the  common  belief  that  coal 
could  be  most  economically  cut  and  blasted  by  using  a  depth 
of  cut  equal  to  the  height  of  seam.  This  erroneous  idea  fre- 
quently resulted  in  blasting  down  more  coal  than  could  be 
loaded  in  one  day,  and  was  the  cause  of  allotting  more  than  one 
room  to  a  miner.  That  the  height  of  seam  does  not  bear  any 
direct  relation  to  economical  cutting  or  blasting  was  demon- 
strated by  the  United  States  Coal  &  Coke  Co.  at  Gary,  W.  Va., 
working  with  a  Sullivan  shortwall  machine,  and  it  has  been 
found  that  miners  are  pleased  to  work  two  or  more  to  a  room, 
provided  their  earnings  are  as  great  as  when  they  work  in 
rooms  by  themselves. 

Probably  the  most  marked  results  in  devising  more  eco- 
nomical mining  methods  has  been  achieved  by  the  officials  of 
the  above-mentioned  company.  They  have  realized  the  objec- 
tions to  the  mining  methods  outlined  above  and  applied  them- 
selves to  working  out  a  plan  which  would  be  simple,  direct  and 
efficient.  They  accepted  it  as  axiomatic  that  any  change  in  the 
prevailing  plans  of  mining  must  be  beneficial  to  the  property 
owner,  operator  and  miner  alike,  for  any  change  that  would 
benefit  one  or  more  of  the  interested  parties  at  the  expense  of 
the  others  would  not  last. 

In  this  study  difficulty  was  experienced  because  of  the  entire 
lack  of  systematized  knowledge  as  to  the  proper  relative  rate  of 
advance  of  room  to  retreat  of  pillar,  the  most  economical  width 
of  room,  and  in  fact  what  might  be  considered  100  per  cent 
efficiency  for  any  man,  animal  or  machine  about  the  mines.  In 
order  to  determine  these  data,  which  were  absolutely  essential 
to  an  intelligent  solution  of  the  problem,  a  series  of  time  studies 
was  instituted  and  extended  over  a  period  of  weeks,  covering 
all  the  motions  that  make  up  certain  underground  operations 
that  have  to  do  with  getting  the  coal  from  the  working  face 
to  the  railroad  car.  Thousands  of  observations  were  taken, 
properly  checked,  tabulated,  collated,  and  used  as  a  basis  for 
a  method  of  procedure,  which  has  been  put  to  the  rigid  test  of 
practical  use  with  remarkably  good  results. 

This  method  of  procedure  has  for  its  object  the  maximum 
degree  of  safety,  sanitation  and  opportunity  to  the  miner  and 
of  security  to  the  property  owner,  while  at  the  same  time  offer- 
ing the  greatest  advantage  to  the  operator.  It  combines  a 


MINING  COSTS  67 

maximum  degree  of  concentration  with  a  minimum  expenditure 
for  labor,  material  and  equipment,  in  such  a  manner  that  these 
quantities  bear  a  constant  relation  to  the  output.  Its  use  has 
resulted  in  a  marked  reduction  in  fatalities,  increased  earnings 
to  the  miners,  decreased  costs  per  ton  for  labor,  material,  equip- 
ment, and  capital,  and  the  recovery  of  practically  all  the  coal 
in  the  seam. 

At  Gary,  W.  Va.,  mules  are  used  for  gathering,  and  as  a 
result  of  concentration  their  efficiency  has,  in  some  instances, 
been  increased  to  over  200  per  cent.  At  one  of  the  mines,  fewer 
day  laborers  are  employed  underground  than  are  employed  about 
the  tipple. 

For  the  purpose  of  comparing  the  results  obtained  under 
this  method  with  those  from  the  several  methods  of  procedure 
in  the  panel  system,  the  method  has  been  applied  to  the  property 
and  the  problem  under  consideration.  Fig.  31  shows  the  arrange- 
ment of  the  workings  at  the  time  the  desired  maximum  out- 
put is  reached.  It  also  shows  the  details  of  the  method  of  pro- 
cedure; the  other  data  desired  are  given  in  Table  II.  Fig.  32 
shows  the  tonnage  curve  and,  for  comparison,  the  total  tonnage 
curve  from  the  unit  entry  shows  that  the  tonnage  rises  rapidly 
curve,  from  the  unit  entry  shows  that  the  tonnage  rises  rapidly 
until  the  maximum  is  reached  and  then  continues  indefinitely  at 
that  rate.  By  using  available  data,  the  proper  length  of  room, 
angle  of  breakline  and  angle  of  advancing  faces  may  be  pre- 
determined, so  that  the  total  daily  tonnage  from  the  entry  is 
the  multiple  of  the  tons  that  can  be  hauled  by  a  mule  or  motor ; 
thus,  the  mules  or  motors  are  always  working  at  the  maximum 
efficiency.  It  is  equally  true  that  when  the  workings  have 
advanced  for  a  short  distance,  after  reaching  the  maximum  ton- 
nage from  the  entries,  the  estimated  minimum  number  of  day 
laborers  required,  may  readily  be  confirmed,  and  once  the  entry 
reaches  its  maximum  tonnage,  and  the  quantities  of  labor, 
material  and  equipment  have  been  accurately  determined,  these 
quantities  remain  constant  throughout  the  entire  extent  of  the 
entry,  which  may  be  as  great  as  the  property  is  long. 

The  room  space  occupied  per  miner  is  less  than  in  any  of 
the  other  methods  now  in  effect,  which  is  an  index  of  the  rela- 
tive degree  of  safety  a  miner  obtains  for  a  given  expenditure 
of  time  and  energy.  The  excellent  manner  in  which  the  rooms 


68 


COAL  MINING  COSTS 


£•8  i  S^^S*  5  C  6*  «  w*  *  ""•  2>V 

I     II  !  I  II 


A, 

^c^  — «\V'V^  »•§  -5-t,  5  •     ^ 


MINING  COSTS  69 

are  timbered,  shown  in  Fig.  31,  is  the  minimum  required ;  where 
the  mine  foreman  or  miner  has  reason  to  believe  that  additional 
timber  is  required  to  make  the  place  safe,  the  miner  must  place 
additional  timber  before  doing  anything  else. 

As  the  entries  advance,  all  rooms  are  driven  and  robbed 
immediately  upon  their  completion,  and  rooms  are  opened  up 
only  fast  enough  to  provide  for  the  uninterrupted  advance  of 


3000 


10        15        20       25        30      35       "JO       45        50       55       60      65 

Months  at  25  Days  per  Month 

Flo.  32. — Curve  showing  rate  of  development  to  the  desired  output,  under 
the  method  of  procedure,  sketch  F,  Fig.  22,  and  the  advancing  plan  of 
mining,  Fig.  31. 

the  robbing.  Thus  no  barrier  pillars  are  required,  for  the  virgin 
coal  protects  the  workings  on  three  sides  and  the  weight  of  the 
roof  is  resting  on  the  bottom  of  the  robbing.  If  a  disturbed 
area  of  coal  is  encountered,  or  for  some  reason  it  is  desired  to 
discontinue  the  panel,  a  barrier  pillar  may  be  introduced  at  any 
time  exactly  where  it  is  needed  and  the  entries  continued  for  the 
purpose  of  exploration. 

In  order  that   the   different   methods   of  mining    may    be 
readily  compared,  Fig.  33,  showing  the  relative  amount  of  labor, 


70 


COAL  MINING  COSTS 


material,  and  equipment  required  to  produce  the  tonnage  desired 
from  the  property  shown  in  Fig.  25 ;  also  the  acreage  of  stand- 
ing pillars,  the  relative  cost  of  production,  and  the  estimated 
percentage  of  recovery. 

Any  method  of  procedure  that  does  not  provide  for  the 
removal  of  pillars  immediately  on  the  completion  of  a  room 
is  fundamentally  wrong,  because  it  involves  long-standing  pillars 
open  to  the  unfavorable  influence  of  atmospheric  agencies  and 
other  forces  of  nature ;  the  duplication  of  track  work ;  the  clean- 


Fferiod  Required  Day  Mining 

to  Reach  the  Laborers  Machines 


Dciired  Output  Required 


Required 


Mine 
Cars 

Required 


Mules 
Required 


I 


Motorii  Mam  EntmTrack       Crois  Mam  Entry, 

Required  and  Trofley  Trackand  Trolley 

Wire  Required  W.re  Required 


FIG.  33. — Comparison  of  the  amount  of  labor,  material  and  property 
required  when  following  the  methods  of  procedure  shown  in  Fig.  25  and 
the  relative  cost  per  ton,  the  recovery,  and  the  period  of  time  required. 

ing  up  of  many  slate  falls  that  might  otherwise  have  been 
avoided;  and  the  scattering  of  workings,  all  of  which  increase 
the  cost  per  ton  for  labor,  material,  and  equipment,  and  cause 
the  pillar  coal  to  be  badly  disintegrated  and  low  in  domestic 
lump  sizes. 

It  sometimes  happens  in  practice,  however,  that  fundamentals 
must  be  sacrificed  to  adapt  the  method  to  peculiar  conditions 
encountered,  often  resulting  in  lack  of  concentration  and  large 
areas  of  standing  pillars.  Where  considerable  tonnage  is  desired 
and  a  new  property  is  being  opened,  skilled  miners,  experienced 
in  robbing  pillars,  are  hard  to  get  and  frequently  the  officials, 


MINING  COSTS 


71 


mine  foreman,  and  underbosses  are  not  experienced.  In  order 
to  keep  up  the  tonnage  under  these  conditions, .  the  workings 
must  necessarily  become  distantly  separated,  because  coal  can 
only  be  obtained  from  room  workings.  It  frequently  happens 
also  that  the  rates  for  mining  pillar  coal  and  room  coal  are  not 
properly  adjusted,  so  that  the  men  can  earn  more  in  room  work 
than  in  pillar  work,  naturally  causing  the  pillars  to  lag  behind, 
and  requiring  the  introduction  of  barrier  pillars  to  safeguard 
against  squeezes;  these  barriers  in  turn  cause  a  further  separa- 
tion of  the  workings,  and  a  decrease  in  the  efficiency  of  labor, 


Room  Entry  and  Roomsand 

Room  Entry  Room  Track 

Track  Required 


Recovery 


FIG.  34.— Equipment  required  to  produce  an  output  of  2800  tons  per  day 
from  the  different  plans  of  mining  outlined.  Also  the  acreage  of  standing 
pillars  to  reach  the  output. 

material,  and  equipment.  The  natural  impulse  of  the  mine  fore- 
man, therefore,  is  to  open  up  more  rooms  in  advance  of  the 
robbing  in  order  to  increase  the  efficiency  to  something  like  a 
proper  standard. 

For  these  reasons  the  territory  for  a  given  output  during 
the  development  period  should  be  as  isolated  as  possible,  and  no 
greater  in  extent  than  is  practicable.  After  the  development 
period  is  passed  and  the  organization  perfected,  there  is  no  good 
reason  why  a  mine  operation  should  not  be  conducted  with  much 
the  same  regularity  as  a  blast  furnace  or  an  industrial  railroad. 

The  fallacy  that  the  average  miner  will  load  only  so  much 


72 


COAL  MINING  COSTS 


coal  and  no  more  has  long  since  been  exploded,  and  it  is  a 
matter  of  every-day  observation  that  the  miners  are  pleased 
to  load  coal  if  the  mine  cars  are  given  to  them  with  some 
degree  of  regularity  and  with  some  relation  to  the  time  required 
to  load  a  car.  When  one  considers  that  a  coke  loader,  work- 
ing under  the  heat  of  the  sun,  and  of  the  coke  ovens,  will 


FIG.  35. — Standard  plan  of  mine  development  adopted  by  the  Pocahontas 

Coal  &  Coke  Co. 

load  from  35  to  40  tons  as  an  ordinary  day's  work,  there  is  no 
reason  why  a  miner  working  under  so  much  more  favorable 
circumstances  should  not  load  at  the  same  rate.  In  this  con- 
nection the  following  observations  that  have  to  do  with  load- 
ing coal  underground  are  interesting. 

These  figures  show  that  less  than  47  per  cent  of  the  time 
spent  underground  was  consumed  in  loading  coal  and  over 
12  per  cent  of  the  time  was  lost  waiting  for  the  empty  mine 


MINING  COSTS 


73 


cars.  It  may  be  stated  further  that  these  men  were  loading 
at  the  rate  of  35  tons  per  day  of  8  hr.,  and  actually  did  load 
at  the  rate  of  16  tons  per  man  per  day  per  year. 

Methods  of  working  in  the  Pocahontas  field.— The  entire 
Pocahontas  field  proper  is  practically  all  leased  out  on  royalty 


FIG.  36. — Double-entry  system  of  mining  used  by  the  Upland  Coal  &  Coke  Co. 

by  two  large  holding  companies,  the  Pocahontas  Coal  &  Coke 
Co.  and  the  Crozer  Land  Association.     Under  the  lease  con- 
tracts, the  holding  companies  have  reserved  the  right  to  define 
the  method  of  working,  and  the  result  has  been  satisfactory. 
A  standard  plan  of  mining,  by  Thomas  H.  Clagett,  chief 


74  COAL  MINING  COSTS 

engineer  of  the  Pocahontas  Coal  &  Coke  Co.,  is  shown  in  Fig.  35 ; 
this  is  largely  followed  by  their  lessees,  although  in  instances 
materially  modified,  due  to  local  conditions.  This  large  hold- 
ing company  owning  or  controlling  some  275,000  acres  of 
Pocahontas  coal,  has  in  active  operation  some  45  leases  (in 
1913),  covering  about  145,000  acres. 

One  of  the  special  advantages  of  the  system  of  mining 
adopted  by  the  Pocahontas  Coal  &  Coke  Co.,  is  the  relatively 
quick  recovery  of  the  pillars,  and  the  panels  are  so  driven 
that  the  rooms  and  all  entries  split  the  pitch;  thus  if  the 
maximum  pitch  is  3  per  cent,  then  the  maximum  for  the  work- 
ing will  not  exceed  2  per  cent  and  may  be  even  slightly  less. 
Further  particulars  of  this  system  appear  on  page  77. 

The  method  of  working  adopted  by  the  Upland  Coal  £ 
Coke  Co.,  on  one  of  the  Crozer  leases,  is  shown  in  Fig.  36. 
Were  the  crop  line  shown  on  this  plan  it  would  be  evident 
that  the  break  line  is  carried  in  from  the  crop  and  does  not 
involve,  strictly  speaking,  breaks  in  the  solid.  There  may  be 
several  " lifts"  where  the  width  of  the  lease  is  too  great  to 
admit  of  one  lift  only,  as  shown.  This  plan  of  mining  was 
evolved  from  a  number  of  years  of  revisions  and  has  been 
found  satisfactory  under  all  conditions.  The  main  entry  is 
to  be  driven  as  near  the  line  of  strike  as  possible,  in  order 
that  the  reverse  grade  against  the  loads  may  be  negligible. 
If  the  cross  entries  are  turned  off  at  more  than  90  deg.  from 
the  main,  and  the  rooms  are  less  than  90  deg.  of  the  cross 
entries,  grades  in  favor  of  the  loads  may  be  obtained. 

The  Pocahontas  Consolidated  Collieries  Co.'s  Angle  col- 
liery, as  of  July,  1912,  is  shown  in  Fig.  37.  Soon  after  taking 
up  the  work  of  the  old  Norfolk  Coal  &  Coke  Co.  (which  was 
essentially  the  nucleus  of  the  Pocahontas  Consolidated  Col- 
lieries Co.)  in  1904,  the  work  of  revising  the  systems  of  mining 
was  taken  up  in  detail.  One  of  the  first  improvements  adopted 
was  the  introduction  of  the  multiple  air-course  system. 

Further  interesting  particulars  regarding  mining  methods 
in  this  field  were  given  in  a  paper  presented  before  the  West 
Virginia  Coal  Mining  Institute  in  1913,  from  which  the  follow- 
ing has  been  excerpted. 

It  is  bad  practice  to  drive  up  a  room  and  allow  the  pillars 
to  stand  in  the  expectation  of  drawing  them  later.  A  better 


MINING  COSTS 


75 


76 


COAL  MINING  COSTS 


method  is  to  start  to  "stump  off"  the  pillars  as  soon  as  the 
room  is  completed  by  cutting  across  the  rib  as  at  A,  in  Fig.  38. 
The  track  should  be  laid  so  that  cars  can  be  placed  con- 
veniently for  loading  out  both  the  coal  in  the  cut  and  also  that 
mined  when  the  stump  is  drawn  back. 

In  this  method,  the  miner  is  always  protected  by  solid  coal 
and  the  losses  are  reduced  to  a  minimum.  Room  No.  6  shows 
the  pillar  and  the  heading  stumps  completely  removed;  room 
No.  5,  a  pocket  just  starting  in ;  room  No.  4,  a  pocket  finished 
and  a  stump  partly  drawn  back.  Room  No.  3  shows  the  pocket 
finished  and  work  just  starting  on  the  stump ;  room  No.  2  shows 
the  pocket  being  driven,  followed  by  a  second  pocket,  which 


FIG.  38. — Method  of  splitting  the  pillars  used  in  the  Pocahontas  Field. 

is  only  extended  as  far  as  a  man  can  conveniently  load  the 
coal  without  a  track  turn,  in  order  to  avoid  the  necessity  for 
frogs  and  switch  points.  Room  No.  1  shows  the  pocket  just 
starting. 

The  width  of  this  pocket  and  the  thickness  of  the  stump 
depend  largely  on  the  nature  of  the  roof  and  the  mine  equip- 
ment. With  poor  roof,  which  falls  unexpectedly  or  within 
a  few  hours  after  the  removal  of  the  coal,  the  thickness  of  the 
stump  should  be  such  that  a  miner  can  reach  all  the  coal 
safely  and  easily  without  venturing  too  far  beyond  the  rib 
line  of  the  pocket.  If  the  roof  is  good  and  does  not  fall  soon 
after  the  removal  of  the  stump  the  thickness  of  the  small 
pillar  may  be  increased  and  the  number  of  track  turns  required 
per  pillar  may  be  reduced. 


MINING  COSTS  77 

In  the  application  of  mining  machines  to  the  robbing  of 
pillars,  the  distance  between  the  centers  of  pockets  should  be 
such  that  the  thickness  of  stump  left  will  form,  under  bad 
roof,  one  machine  cut,  or  under  good  roof,  two  cuts  of  the 
machine. 

The  more  common  practice  where  the  roof  falls  soon  after 
the  extraction  of  the  stump  is  to  leave  a  small  shell  of  coal 
to  protect  the  miner  from  the  gob  and  also  prevent  him  from 
loading  fine  slate  into  the  car  of  coal.  This  results  in  a  loss 
of  coal  that  can  be  avoided  at  an  expense  for  timber,  which, 
under  ordinary  circumstances,  is  less  than  the  value  of  the 
coal. 

A  practice  which  has  been  advocated  and  proved  success- 
ful, is  to  place  a  row  of  props  on  the  lower  rib  of  the  pocket, 
before  the  removal  of  the  pillar  stump  has  begun.  When  the 
next  pocket  to  the  outby  is  driven,  it  will  be  found  that  prac- 
tically the  entire  stump  may  be  loaded  out  without  any  admix- 
ture of  gob  and  a  greater  percentage  of  lump  coal  will  be 
obtained.  This  precautionary  row  of  timbers  is  especially 
desirable  where  machines  are  used  on  the  pillars. 

The  roof  over  a  robbing  line  exceeding  2400  ft.  in  length 
sometimes  begins  to  sag  in  the  middle  and  renders  the  removal 
of  the  pillars  in  that  immediate  section  difficult. 

The  breakline  should  be  kept  as  uniform  as  possible  at 
all  times.  A  method  in  practical  every-day  use,  which  is  to 
be  recommended,  is  as  shown  in  Fig.  39. 

The  engineers,  as  they  take  their  monthly  measurements, 
mark  the  pocket  centers  on  the  robbing  rib  of  the  room,  and 
the  foremen  are  required  to  drive  their  crosscuts  on  the  line 
of  a  pocket  as  at  A.  The  object  in  keeping  the  breakline 
uniform  is  to  insure  against  pillars  extending  back  into  the 
gob  and  acting  as  a  fulcrum,  or  the  knife-edge  of  a  scale  beam, 
upon  which  the  roof  teeters.  This  almost  invariably  causes 
additional  timber  expense  and  sometimes  losses  of  coal,  both 
of  which  could  have  been  avoided  had  the  break  line  been 
kept  uniform. 

The  essential  features  of  the  Pocahontas  Coal  &  Coke  Co.'s 
plan  of  mining,  shown  in  Fig.  35  are :  Provisions  for  tonnage 
during  the  development  period;  provision  for  meeting  the 
market  demand ;  large  barrier  pillars,  insuring  against  squeezes 


78 


COAL  MINING  COSTS 


and  rendering  impossible  the  destruction  of  coal  over  an 
extended  area ;  four-entry  system  for  all  extensive  main  entries, 
using  two  as  intakes  and  two  as  returns  with  breakthroughs 
between  only  at  points  where  the  cross  entries  turn  off,  render- 
ing unnecessary  the  building  of  expensive  masonry  brattices 
every  80  ft.  and  insuring  the  maximum  quantity  of  air  for 
ventilation  at  a  minimum  cost  for  brattices  and  ventilating 
power;  cross  entries  with  narrow  chain  pillars,  permitting 
the  rapid  advance  of  the  entry. 

In  general  all  robbing  must  be  done  retreating  with  rooms 
driven  after  the  entry  is  nearing  completion,  insuring  against 
slate  falls  and  rendering  possible  the  extraction  of  all  of  the 


FIG.  39. — Locating  crosscuts  so  as  not  to  interfere  with  lifts  from  pillar. 

coal  in  one  operation,  thus  combining  first  development  and 
robbing. 

The  depth  and  number  of  rooms  on  an  entry  vary  greatly 
at  different  mines,  depending  on  local  conditions  of  the  seam; 
whether  the  haulage  is  by  mule  or  gathering  motor,  whether 
the  undercutting  is  performed  by  pick  or  machine,  and  not 
infrequently  on  the  personal  equation  of  the  mine  executive, 
for  sometimes  the  manager  of  a  plant  will  contend  that  he 
obtains  the  best  results  when  he  drives  25  rooms,  500  ft,  deep 
to  the  entry,  and  another  manager  working  on  an  adjoining 
property  under  identically  the  same  physical  conditions  and 
with  the  same  type  of  equipment,  not  1000  ft.  away,  will  say 
that  he  gets  the  best  results  when  his  rooms  do  not  exceed 
300  ft.  in  depth  and  when  there  are  only  15  rooms  to  the 


MINING  COSTS 


79 


entry.  The  better  policy  is  to  encourage  individual  initiative 
and  allow  freely  such  modifications  in  any  plan  of  mining  as 
may  be  desired,  provided  that  the  modified  plan  embodies  all 
the  principles  of  modern  methods  and  sound  mining  practice. 

In  the  successful  operation  of  any  mine  some  general  scheme 
of  mining  must  be  agreed  upon,  subscribed  to  by  all  parties 
in  any  way  concerned  with  the  matter,  including  the  land 
owner,  if  the  property  is  a  leased  one.  Then  no  omissions  in, 
additions  to,  or  deviations  from  that  plan  of  mining  should  be 
permitted  without  the  written  consent  of  the  lessee  and  lessor. 

RECOVERY  OF  COAL  IN  MINES  OF  POCAHONTAS  COAL  &  COKE  Co. 


Plant 

Thick- 
ness of 
Seam 
in 
Feet 

Acres 
of 
Entry- 
mined 

Acres 
of 
Rooms 
Mined 

Acres  - 
of 
Pillars 
Mined 

Total 
Acres 
Mined 

Total 
Tonnage 
Mined 

Tons 
Mined 
per 
Acre 

Theo- 
retical 
Tons 
per 
Acre 

Per- 
centage 
of  Re- 
covery 

Propor- 
tion of 
Seam 
Re- 
jected 

1 

6.15 

3.06 

4.57 

11.03 

18.66 

165,254 

8,856 

9,922 

89.3 

0.24 

2 

5.65 

4.40 

4.80 

14.80 

24.00 

188,391 

8,185 

9,115 

89.79 

0.22 

3 

5.16 

2.68 

6.52 

15.80 

25.00 

180,386 

7,215 

8,325 

86.6 

0.22 

4 

4.42 

5.88 

8.65 

13.09 

27.62 

192,437 

6,960 

7,131 

97.6 

0.23 

5 

5.94 

7.00 

10.09 

19.20 

36.29 

334,005 

9,203 

9,582 

96.0 

0.22 

6 

4.32 

2.11 

3.64 

9.20 

15.04 

94,427 

6,278 

6,969 

90.0 

0.31 

7 

5.34 

3.31 

6.34 

0.00 

9.65 

83,000 

8,601 

8,614 

99.8 

0.20 

8 

5.42 

3.72 

6.06 

9.72 

19.50 

144,769 

8,181 

8,777 

93.2 

0.20 

9 

4.65 

8.10 

16.80 

2.34 

27.24 

201,044 

7,380 

7,534 

98.0 

0.18 

10 

8.03 

5.20 

8.47 

10.09 

23.76 

262,975 

11,068 

12,923 

85.6 

0.23 

After  the  general  plan  of  mining  has  been  decided  and 
operations  begun,  its  success  or  failure  will  depend  largely 
upon  the  degree  of  watchfulness  exercised.  The  mine  should 
be  accurately  surveyed  and  mapped  at  least  once  every  90 
days.  Frequent  inspections  should  be  made  of  the  mine,  minute 
attention  being  given  to  the  conditions  of  the  working  faces 
and  the  robbing  line.  At  least  once  a  year  the  theoretical 
yield  of  the  property  should  be  balanced  against  the  actual 
tonnage  delivered  at  the  tipple.  Accurate  and  complete  records 
should  be  kept  of  the  number  of  acres  of  entries,  and  rooms 
driven  and  pillars  drawn  each  year  and  of  both  the  percentage 
of  recovery  per  acre  and  the  state  of  exhaustion  of  the  prop- 
erty. 

That  the  above  method  of  mining  will  yield  the  maximum 
recovery  is  indicated  in  the  accompanying  table,  the  figures  in 


80  COAL  MINING  COSTS 

which  are  typical  of  the  results  obtained  by  the  Pocahontas 
Coal  &  Coke  Co.,  which  are  probably  unexcelled  anywhere. 
In  this  connection  it  should  also  be  noted  that  the  percentages 
of  recovery  are  based  on  the  total  seam,  including  the  portion 
rejected. 

The  lower  percentages  of  recovery  in  the  table  result  from 
the  fact  that  in  some  instances,  pillars  were  being  robbed  that 
had  been  standing  for  many  years.  In  the  mines  of  the  United 
States  Coal  &  Coke  Co.,  at  and  near  Gary,  W.  Va.,  where 
all  the  work  has  been  opened  in  recent  years,  the  average 
percentage  of  recovery  per  acre  since  the  beginning  of  opera- 
tions, has  been  better  than  95  per  cent,  and  of  the  area  mined, 
over  one-third  has  been  final  mining.  The  cost  of  production 
of  room  and  entry  coal  by  this  method  is  the  same  as  in  other 
methods  of  mining,  while  pillar  coal  is  produced  at  less  expense 
than  is  incurred  in  other  methods. 

Most  operating  companies  have  a  statement  showing  the 
revenue  derived  from  operations  per  ton  of  coal  mined  based 
on  the  net  receipts  from  operations  divided  by  the  tonnage. 
By  placing  a  value,  which  could  be  closely  approximated,  on 
the  recoverable  coal  lost,  and  adding  it  to  the  net  receipts  a 
figure  could  be  obtained  showing  what  revenue  would  have 
been  derived  from  the  operations  had  the  coal  been  mined 
without  unnecessary  waste.  Dividing  this  by  the  tonnage  a 
figure  could  be  obtained  for  the  statement  which  would  show 
the  profit  that  would  have  been  derived  per  ton  produced  if 
the  coal  in  the  seam  had  been  worked  by  the  most  conservative 
methods. 

Statements  of  the  above  nature  have  a  further  value  from 
a  financial  point  of  view  for  if  it  can  be  shown  to  a  bonding 
concern  that  a  property  contains,  let  us  say  $500,000  worth  of 
coal  in  the  ground,  90  per  cent  or  more  of  which  will  be  mined, 
it  is  certain  that  a  greater  asset  value  will  be  placed  on  the 
property  than  would  be  credited  to  it  if  the  engineers  of  the 
bonding  house  report  that  under  the  methods  of  mining  pur- 
sued, only  50  per  cent  of  the  coal  in  the  ground  will  be  mined 
and  the  rest  irretrievably  lost. 

Connellsville  Systems. — The  system  of  mining  practiced  by 
the  H.  C.  Frick  Coke  Co.  in  the  Connellsville  region  as  described 
in  a  paper  presented  before  the  Engineers  Society  of  Western 


MINING  COSTS  81 

Pennsylvania  in  1916,  is  the  application  of  shortwall  mining 
machines  to  the  extraction  of  rib  coal.  The  two  salient  factors 
effecting  this  result  were,  first,  the  effort  to  reduce  accidents 
and  second,  the  desire  to  obtain  an  increased  output  of  coal 
per  man  per  day. 

It  has  long  been  realized  that  the  more  intense  the  super- 
vision of  working  places  and  workmen  the  less  liability  there 
is  to  accident.  In  order  to  obtain  the  desired  supervision 
without  making  the  cost  prohibitive,  it  was  seen  that  the  time 
spent  by  mine  officials  in  traveling  from  working  place  to 
working  place  must  be  reduced  to  the  minimum  and  the  time 
actually  spent  in  working  places  increased  to  the  maximum. 
To  obtain  this  result  the  working  places  were  concentrated 
gradually,  and  it  was  soon  found  that,  under  the  old  method 
of  pick  mining,  a  limit  was  quickly  reached,  and  it  was  realized 
that  to  obtain  the  desired  intensive  supervision  it  was  necessary 
to  decrease  the  number  of  working  places  and  workmen.  This 
could  only  be  accomplished  by  an  increased  production  from 
each  miner  and  a  consequent  reduction  in  the  number  of  work- 
ing places  without  affecting  the  total  output  of  the  mine. 

On  account  of  the  conditions  in  the  Connellsville  region, 
where  it  is  necessary  to  drive  narrow  headings,  narrow  rooms 
and  have  large  room  centers,  it  was  found  that  machines  in 
the  narrow  work  would  not  accomplish  the  result  since  the 
bulk  of  the  coal  comes  from  rib  extraction. 

The  use  of  electrically  driven  mining  machines  and  the 
blasting  of  coal  on  the  very  rib  line  itself  requires  a  system 
of  ventilation  that  will  insure  gob  gas  being  found  only  on  the 
return.  Such  a  system  of  ventilation  necessitates  ample, 
reliable  fan  equipment,  airways  of  sufficient  size  and  number, 
a  generous  provision  of  upcast  openings,  wise  coursing  of  the 
air  current  and  the  existence  of  numerous  bleeders  from  every 
gob  into  a  return  airway.  It  also  demands  the  elimination 
of  danger  from  dust  by  keeping  it  sprinkled  and  removing  it 
before  any  dangerous  accumulations  are  found.  In  addition  it 
necessitates  the  use  of  permissible  explosives  and  these  only 
in  the  hands  of  selected  competent  shotfirers. 

It  has  been  proved  that  working  places  cannot  be  concen- 
trated to  as  great  an  extent  by  any  system  yet  tried  in  the 
Connellsville  region  as  by  the  use  of  the  H.  C.  Frick  Coke  Co.  's 


82  COAL  MINING  COSTS 

system  of  machine  mining  in  rib  coal.  On  account  of  the 
intense  concentration  of  working  places  and  the  output  that 
is  obtained  it  is  necessary  that  the  haulage  arrangements  and 
equipment  be  perfected  beyond  anything  that  had  previously 
been  necessary,  and  the  transportation  of  coal  cannot  well  be 
handled  except  by  the  use  of  electric  gathering  locomotives. 

The  general  plan  by  simple  modifications  can  be  made  to 
suit  all  conditions;  depth  of  cover;  presence  or  absence  of 
drawslate ;  nature  of  coal,  and  the  nature  of  bottom  and  roof. 
This. is  divided  into  what  we  know  as  maximum,  medium  and 
minimum  plans. 

The  maximum  plan  is  applicable  where  thickness  of  over- 
lying cover  does  not  exceed  125  ft.  and  where  the  coal  is  hard 
and  the  general  physical  condition  of  roof  and  bottom  is  good. 

The  medium  plan  is  applied  where  the  cover  does  not  exceed 
250  ft.  with  the  same  physical  conditions  of  coal  and  bottom 
and  roof  as  obtain  under  the  maximum  plan. 

The  minimum  plan  may  be  applied  to  coal  underlying  any 
depth  or  thickness  of  cover,  and  whether  or  not  the  coal  is 
hard  or  soft  and  the  physical  condition  of  roof  and  bottom 
good  or  bad,  provided,  of  course,  that  mining  machines  in  any 
form  can  be  used. 

The  H.  C.  Frick  Coke  Co.  has  always  worked  its  mines 
according  to  a  projection,  carefully  prepared,  for  the  field  of 
coal  to  be  worked  before  actual  excavations  have  been  started. 
In  this  plan  of  concentrated  mining  it  has  been  found  of  great 
advantage  to  supplement  these  general  projections  with  a 
schedule,  prepared  on  a  scale  of  20  ft.  to  the  inch,  showing  in 
detail  the  daily  operations. 

It  should  be  understood  that  in  the  shortwall  plan  of  min- 
ing the  development  is  made  on  the  face  and  the  butt  of  the 
coal.  After  it  has  been  determined  as  to  what  plan  is  to  be 
followed  for  a  given  tonnage,  the  mining  section  is  projected 
and  developed  and  a  fracture  line  established. 

Let  us  first  consider  the  minimum  plan  of  extraction.  The 
main  haulage  headings  are  driven  on  the  face  as  are  also  the 
return  airways  while  the  producing  headings  are  on  the  butt. 
Off  these  producing  headings  main  face  rooms  are  turned, 
generally  on  112-ft.  centers.  From  these  main  face  rooms,  butt 
rooms  are  driven  on  25-ft.  centers.  As  the  main  face  rooms 


MINING  COSTS  83 

advance  the  necks  of  the  butt  rooms  to  be  driven  are  excavated 
to  a  depth  of  three  machine  cuts.  After  this  main  face  room 
has  been  advanced  50  ft.  there  are  available  two  places  for  the 
machine  to  cut  that  will  yield  40  tons,  and  when  it  advances 
to  a  point  where  the  first  crosscut  is  turned  off,  there  are  three 
places  to  cut  in  each  main  face  room,  yielding  60  tons.  This 
main  room  may  continue  to  the  end  of  the  section  or  to  the 
end  of  the  coal  field,  turning  butts  or  producing  headings  off 
at  projected  distances.  The  main  face  rooms  being  driven  on 
112-ft.  centers  and  12  ft.  wide  leave  a  pillar  100  ft.  in  thickness 
between  the  rooms.  This  pillar  is  considered  ample  to  support 
any  thickness  of  cover  with  a  floor  or  bottom  under  the  coal 
seam  of  any  nature  that  may  be  found  in  the  Connellsville 
region. 

On  this  minimum  plan  of  extraction,  where  main  rooms 
are  advanced  sufficiently  far  to  begin  the  extraction  of  main 
face  room  pillars,  the  butt  rooms  are  advanced  in  succession 
so  that  each  room  is  50  ft.  behind  the  one  next  preceding. 
This  plan  provides  for  a  tonnage  output  from  three  working 
places — two  butt  rooms  advancing  furnish  40  tons  and  one  butt 
rib  retreating  provides  an  additional  40  tons,  or  a  total  of 
80  tons  of  coal  while  retreating;  and  the  main  face  room 
advancing  is  yielding  60  tons,  or  a  total  of  140  tons  of  coal 
from  one  main  face  room  properly  prepared  and  developed  on 
this  minimum  plan  of  production.  A  sketch  of  this  method  is 
shown  in  Fig.  40. 

Along  the  same  lines  the  medium  plan  will  not  yield  any 
greater  tonnage  from  the  advancing  main  rooms,  but  on  the 
retreat  the  butt  rooms  are  so  driven  as  to  maintain  each  face 
30  ft.  behind  that  of  the  preceding  room.  This  allows  three 
butt  rooms  to  advance  at  a  time,  producing  60  tons,  and  neces- 
sitates two  butt  ribs  retreating  at  the  same  time,  giving  a 
production  of  80  tons,  or  a  total  from  the  butt  rooms  of  140 
tons.  This  with  the  production  of  60  tons  from  the  advancing 
main  room  totals  200  tons  for  each  main  room.  This  method  is 
shown  in  Fig.  41. 

In  the  maximum  plan  the  main  face  rooms  advancing  pro- 
duce 60  tons  while  the  butt  rooms  are  so  driven  that  the  face 
of  one  is  15  ft.  behind  the  face  of  the  preceding  room,  thus 
necessitating  four  advancing  butt  rooms  and  the  simultaneous 


84 


COAL  MINING  COSTS 


.    ung 


MINING  COSTS  85 

withdrawal  of  four  butt  ribs.  The  four  advancing  butt  rooms 
will  produce  80  tons  while  the  four  retreating  butt  rooms  will 
produce  160  tons.  The  sum  of  these,  together  with  the  60 
tons  produced  by  the  advancing  main  room,  gives  a  total 
tonnage  of  300  for  each  main  room.  The  maximum  plan  is 
shown  in  Fig.  42. 

The  work  is  thoroughly  systematized  and  the  routine  can 
be  described  as  follows:  After  the  miner  has  cleaned  up  his 
place  and  the  day's  run  is  completed  the  machine  crew  enters 
and  cuts  the  place  to  a  depth  approximately  7  ft.  The  timber- 
men  follow  the  machine  crew,  resetting  any  posts  that  it  has 
been  necessary  for  the  machine  men  to  remove.  They  post 
up  any  crossbars  that  have  been  notched  in  the  coal  over  the 
machine  cut,  and  generally  put  the  place  in  good  condition, 
following  out  a  prescribed  system  of  timbering.  The  timber- 
men  are  followed  by  the  driller,  who  bores  the  holes  for  blast- 
ing with  an  electrically  operated  power  drill.  The  driller  is 
followed  by  the  shotfirer,  who  charges  the  hole,  tamps  it,  and 
after  his  own  personal  examination  of  conditions,  explodes  the 
charge,  blasting  down  the  coal  ready  for  loading.  After  the 
coal  has  been  blasted  empty  cars  are  placed  by  the  gathering 
locomotives  preparatory  for  the  next  day 's  work,  so  that  when 
the  loader  arrives  at  his  working  place  in  the  morning  it  is  in 
a  safe  condition  and  every  facility  has  been  given  him  to  load 
a  maximum  tonnage.  Especial  pains  are  taken  through  the 
day  to  see  that  wagons  are  changed  as  soon  as  loaded,  thereby 
eliminating  all  unnecessary  loss  of  time  and  allowing  the  men 
to  load  a  maximum  tonnage  in  the  minimum  time. 

Actual  results  obtained  regularly  with  miners  loading  under 
these  conditions  are  18  to  20  tons  per  man  per  shift ;  the  average 
of  all  the  loaders  behind  shortwall  mining  machines  in  all 
mines  of  the  company  for  the  month  of  August,  1916,  was 
approximately  19  tons  per  shift. 

At  mines  where  there  is  a  full  equipment  of  mining  machines 
the  proportion  of  machine-coal  amounts  to  from  80  to  95  per 
cent  of  the  total  output. 

The  concentration  of  work  that  has  been  obtained  by  this 
method  resulted  in  a  decrease  in  the  cost  of  transportation, 
ventilation,  track  work  and  drainage  because  of  the  smaller 
area  in  active  operation.  There  is  also  a  considerable  saving 


86 


COAL  MINING  COSTS 


in  the  amount  money  invested  in  track  and  materials  generally. 

Some  further  interesting  data  on  this  system  appeared 
in  a  paper  presented  before  the  Coal  Mining  Institute  of 
America  from  which  the  following  has  been  excerpted. 

Preparations  for  the  adoption  of  this  system  are  made  well 
in  advance  by  subdividing  the  panels  into  blocks  90  ft.  square 


FIG.  43. — Plan  showing  how  the  Connellsville  system  is  used  at  the  Con- 
tinental Mine  No.  2. 

as  shown  in  Fig.  44.  Double  butt  entries  on  50-ft.  centers, 
10  ft.  in  width  and  1200  ft.  long  are  driven  in  parallel  across 
the  panel  with  cutthrough  connections  every  100  ft.  Other 
and  similar  butt  entries  are  driven  350  ft.  apart,  dividing  the 
panels  into  blocks  350  by  1200  ft.  and  the  chain  pillars  between 
the  double  entries  into  blocks  40  by  90  ft. 

The  350  by  1200-ft.  blocks  are  then  subdivided  into  blocks 
90  ft.  square  by  driving  rooms  10  ft.  wide  and  350  ft,  long 


MINING  COSTS 


87 


at  right  angles  to  the  butt  entries,  the  rooms  being  connected 
by  cutthroughs  at  intervals  of  100  ft.  In  this  manner  a  whole 
panel  can  be  developed  and  prepared  to  a  reasonable  distance 
in  advance  for  the  work  of  concentration  in  quickly  withdraw- 
ing the  pillar  coal. 

Fig.  45  illustrates  the  concentration  method,  showing  a  part 
of  a  section  when  in  full  operation  as  worked  in  the  mines  of 
the  lower  Connellsville  district.  This  shows  the  coal  in  90-ft.- 
square  blocks  and  oblong  pillars  15  ft.  by  90  ft.,  also  the  entries, 


Regulator..,^ 


FIG.  44. — Method  of  laying  out  mine  in  90-ft.  blocks  used  in  the 
Connellsville  region. 

rooms,  crosscuts  and  cutthroughs.  The  section  shown  cross- 
hatched  represents  that  portion  from  which  the  coal  has  all 
been  withdrawn  and  the  overlying  strata  have  subsided,  or 
the  "gob"  section. 

This  section  of  coal  is  developed  by  driving  on  the  right 
hand  side  a  pair  of  butt  entries,  10  ft.  in  width,  500  ft.  in 
length,  on  50  ft.  centers  and  connected  by  cutthroughs  every 
100  ft.  Rooms  10  ft.  wide  and  350  ft.  long,  on  100  ft.  centers, 
are  driven  at  right  angles  to  the  butt  entries,  the  rooms  being 
connected  by  a  straight  line  of  cutthroughs  at  distances  of 
100  ft.  apart  for  ventilation.  Thus,  the  room  pillars  are  divided 
into  3y2  blocks  90  ft.  square,  as  shown  on  the  plan  in  prepara- 


88 


COAL  MINING  COSTS 


tion  for  the  final  operations  of  driving  the  crosscuts  and  the 
withdrawal  of  the  pillar  coal. 

The  selection  of  the  place  to  commence  on  the  pillar  work 
is  important  and  must  be  determined  by  the  persons  directly 
interested  largely  from  the  local  conditions — such  as  the 
inclinations  or  pitch  of  the  coal  bed,  convenience  of  transpor- 
tation or  haulage,  the  size,  area  or  extent  of  the  panel  or 
section  available  for  operations  or  the  required  daily  tonnage. 

Preparations  are  now  completed  for  the  essential  part  of 
the  concentration  work.  It  will  be  noticed  on  the  map  that 
the  pillar  work  at  the  end  of  the  last  room,  which  was  in  the 


FIG.  45. — The  concentration  method  used  in  the  Lower  Connellsville  region. 

upper  left-hand  corner  of  the  plan,  has  been  started.  Each 
room  having  been  divided  into  3y2  blocks,  90  ft.  square,  almost 
three  and  a  half  blocks  from  the  last,  or  No.  5  room  have  been 
worked  out;  nearly  one  and  a  half  blocks  from  No.  4  room; 
one  fourth  of  a  block  from  No.  3  room  and  crosscuts  started 
in  No.  2  room.  This  makes  the  angle  of  the  gob  line  about 
45  deg.  with  the  butt  entry. 

The  plan  shows  that  the  pillar  work  of  each  room  is  75  ft. 
in  advance  of  the  pillar  work  or  gob  fall  of  the  room  follow- 
ing. The  room  pillars  are  kept  in  this  position  for  the  purpose 
of  breaking  the  roof  falls  in  the  advancing  pillars  and  offering 
more  resistance  to  thrust  caused  by  the  breaking  of  the  strata ; 
also  for  the  purpose  of  affording  better  protection  against 
crushing  in  the  pillars  of  the  room  following. 


MINING  COSTS  89 

The  system  so  far  as  this  pair  of  butt  entries  is  concerned 
is  now  in  full  operation.  The  pillar  work  or  gob  line,  however, 
may  be,  and  often  is,  extended  across  several  pairs  of  butt 
entries  (see  Fig.  45),  leaving  an  offset  of  about  75  ft.  as  a 
breaker  at  each  pair.  In  this  manner  the  gob  line  may  be 
extended  clear  across  the  panel  at  an  angle  of  45  deg.  or  even 
at  35  deg. 

The  work  on  this  plan  was  commenced  at  the  top  of  the 
last  room  in  the  block,  which  was  in  the  upper  left-hand  corner 
of  the  map.  New  crosscuts  were  started  at  intervals  of  two 
days,  thus  making  each  crosscut  two  days'  work,  or  about 
12  ft.  in  advance  of  the  one  following,  until  the  crosscut  first 
started  has  been  driven  through  the  90-ft.  block,  for  which 
15  days  were  allowed  at  6  ft.  per  day. 

The  crosscuts  being  10  ft.  in  width  and  turned  off  the  rooms 
at  distances  of  25  ft.  between  centers  makes  the  pillars,  for 
the  final  operation,  15  ft.  in  width  and  90  ft.  in  length.  These 
pillars  are  divided  and  subdivided  by  lines  drawn  across  and 
lengthwise  of  the  pillar.  The  cross  division  lines  divide  the 
pillar  into  three  sections  of  15  X  30  ft.  each — the  amount  of 
coal  to  be  taken  out  in  one  fall  by  three  or  four  men  working 
two  days — which  makes  three  falls  to  each  pillar  in  six  days' 
work,  as  shown. 

Three  of  these  pillars  in  each  room  are  being  worked  at 
the  same  time  and  are  started  at  intervals  of  two  days,  thus 
placing  each  pillar  two  days*  work,  or  30  ft.  in  advance  of 
the  one  following. 

The  proper  time  to  start  the  first  crosscut  at  the  top  of 
the  next  room  in  order  that  the  75-ft.  offset  may  be  maintained 
can  be  ascertained  as  follows :  It  takes  15  days  to  complete  the 
first  crosscut  at  the  top  of  the  last  room,  six  days  to  withdraw 
the  pillar,  two  days  to  finish  the  next  pillar  and  two  more 
days  to  finish  the  next  one,  making  a  total  of  25  days  to  com- 
plete the  withdrawal  of  the  three  pillars  or  cover  a  horizontal 
distance  of  75  ft.  Therefore,  by  starting  the  first  crosscut 
in  the  next  room  10  days  later,  the  offset  of  75  ft.  will  be  main- 
tained as  shown  on  the  plan. 

By  continuing  the  work  on  this  schedule,  leaving  intervals 
of  two  days  between  the  starting  time  of  each  crosscut  and 
10-day  intervals  between  the  starting  time  of  the  crosscut  in 


90  COAL  MINING  COSTS 

the  next  room,  we  shall  have  three  room  pillars  of  90-ft.  thick- 
ness retreating  in  a  diagonal  line  on  each  pair  of  butt  entries 
and  three  of  the  crosscut  pillars  in  each  room  pillar  or  9 
working  places  when  in  full  operation. 

In  estimating  the  amount  of  daily  output  from  this  section, 
there  are  in  operation  15  crosscuts  and  9  pillars.  Allowing  one 
man  in  each  crosscut  and  three  men  to  each  pillar  makes  a  total 
of  42  loaders.  The  coal  being  undercut  to  a  depth  of  6  ft.  by 
the  mining  machines,  the  tonnage  of  each  crosscut  6  ft.  undercut, 
10  ft,  in  width  and  7  ft.  in  height  of  seam,  will  be  6X10X7=420 
cu.  ft.  Allowing  80  Ib.  to  each  cubic  foot  and  2000  Ib.  to  the 
ton,  we  have, 

420X80 

^000- =  16.8  tons, 

and  15  crosscuts  equals  16.8X15  or  252  tons.  Assuming  a  two- 
ton  car,  this  gives  us  252/2=126  loaded  cars. 

From  the  9  pillars  having  6-ft.  depth  of  undercut,  30  ft. 
in  length  of  pillar  and  7  ft.  in  thickness  of  seam,  there  will  be 

6X30X7X80 


2000 


=  50.4  tons, 


and  for  9  pillars, 

50.4X9=453.6  tons,  or  226.8,  2-ton  loaded  cars, 

which  makes  the  total  number  of  tons  from  the  whole  section 

252+453.6  =  705.6  tons,  or  352.8  loaded  cars  from  42  loaders, 

being  an  average  of  8.4  cars  for  each  loader.  This  average  may 
seem  rather  high  until  we  take  into  consideration  the  facilities 
afforded  by  the  concentration  method  and  the  preparations  made 
to  enable  the  miner  or  loader  to  attain  a  high  efficiency. 

During  the  previous  night  all  working  places  are  under- 
cut to  a  depth  of  6  ft.  by  mining  machines.  The  shot  holes 
are  drilled  with  power  drills  by  men  employed  especially  for 
that  purpose  and  are  charged,  tamped  and  fired  by  shotfirers 
using  permissible  explosives,  clay  for  tamping  and  electric 
batteries  for  firing. 

By  this  system  each  miner,  when  he  arrives  at  his  work- 
ing place,  has  about  eight  or  nine  carloads  of  loose  coal,  which 
he  can  at  once  begin  to  load,  provided  no  unusual  difficulties 


MINING  COSTS  91 

arise  to  prevent  him.  He  is  kept  constantly  supplied  with 
empty  cars  by  the  driver,  who  can  also  attain  a  high  efficiency 
by  reason  of  having  the  loaders  within  a  comparatively  small 
area  and  only  a  short  distance  from  the  side  track.  The 
tracks  are  kept  in  good  condition  and  laid  with  steel  rails, 
even  in  the  miner's  places. 

The  trackmen,  timbermen  and  laborers  are  also  enabled 
to  do  more  effective  work,  as  there  is  no  lost  time  in  traveling 
long  distances  from  place  to  place.  For  the  same  reason  much 
better  supervision  can  be  given  by  mine  foreman,  assistant 
mine  foreman  and  firebosses  to  the  workmen  and  working 
places.  They  can  make  frequent  visits  and  keep  in  close  touch 
with  the  workmen  and  other  matters  influencing  the  success 
of  the  operations — such  as  the  machinery,  transportation  or 
haulage,  ventilation,  trackmen  and  laborers,  timbering  and 
timbermen,  miners  and  drivers — and  see  that  defects  interfer- 
ing with  work  or  causing  delays  are  remedied  immediately. 

Comparative  cost  of  mining  different  thickness  of  coal. — 
An  interesting  study  of  the  determination  of  the  minimum 
thickness  of  anthracite  coal  that  can  be  economically  mined 
was  presented  in  a  paper  read  before  the  Engineering  Society 
of  Northeastern  Pennsylvania  in  1914. 

In  deciding  which  beds  are  and  which  are  not  minable, 
we  face,  at  once,  the  question  of  profitable  operation,  and  it 
may  be  conceded  that  other  things  being  equal,  beds  which  are 
6  ft.  and  over  in  thickness  are  more  cheaply  mined  than  those 
which  are  thinner.  If  we  eliminate  all  variations  other  than 
cutting  and  loading  in  making  our  calculations,  the  relative 
cost  of  mining  for  varying  thicknesses  is  a  matter  of  simple 
calculation. 

Let  a  =  allowance  for  rock  in  cents  per  inch  per  yard; 

h  =  normal  required  thickness,  in  inches,  on  which  allowance  is  based; 
x  =  thickness  of  coal,  in  inches,  as  measured  for  allowance; 
xl=uei  thickness  of  coal,  in  inches; 
/= capacity  of  mine  car,  cubic  feet; 
iw= width  of  chamber  in  feet; 

s  =  thickness,  in  inches,  which  will  give  one  mine  car  per  yard  of 
chamber.     Assume  loose  coal  occupies  1|  times  the  volume  of 
an  equal  weight  of  solid; 
c  =  cents  per  car  allowance; 
m  =  mining  price  per  car,  no  allowance. 


92  COAL  MINING  COSTS 


Then 


(h—  x)  a  =  allowance  per  yard  of  chamber; 


/.   s  =  -  =  number  cars  per  yard  of  chamber  width; 
3ws 

(h—x)a    has—asx 


From  this,  for  any  particular  conditions,  the  cost  for  each 
thickness  may  be  calculated,  and  a  curve  constructed  show- 
ing the  relation  between  cost  and  thickness,  as  shown  in  the 
diagram,  Fig.  46. 


Thickness  of  Coal 
2345 


O    10  EO  30  40  50  60  70  80  90  100  110  120  130  140 

Percent  Increase  m  Output 
FIG.  46. — Relation  of  thickness  of  seam  and  output  to  cost  of  coal. 

Unfortunately,  all  the  costs  which  vary  with  the  thickness 
of  bed  are  not  susceptible  to  calculation,  but  in  general  the 
inside  costs  increase  rapidly  with  the  mining  of  thinner  coal, 
and  the  diagram  showing  the  variation  of  cost  with  thickness 
is  believed  to  represent  fairly  the  average  conditions  in  the 
anthracite  fields.  It  was  constructed  by  plotting  a  large  num- 
ber of  actual  costs  and  then  drawing  the  average  curve. 


MINING  COSTS  93 

We  find  ourselves  facing  the  dilemma  whether  we  shall 
mine  the  coal  which  is  profitable  in  itself  or  a  mixture  of  profit- 
able and  unprofitable  coal  which,  through  the  preponderance 
of  the  latter,  will  result  in  a  profitable  output.  Our  first 
thought  would  naturally  be  that  only  the  coal  which  is  actually 
profitable  should  be  mined,  but  when  we  make  a  more  com- 
plete inquiry  we  find  another  condition  in  the  relation  of  out- 
put to  cost.  As  but  approximately  one-third  of  the  inside  cost 
is  expended  in  actual  cutting  and  loading,  and  as  the  greater 
part  of  the  outside  cost  is  but  slightly  dependent  upon  output, 
the  cost  per  ton  will  be  found  to  vary  greatly  with  the  coal 
production,  even  with  a  fixed  unit  cost  for  cutting  and  load- 
ing. How  great  this  variation  may  be  is  indicated  on  the 
diagram,  and  it  is  apparent  that  a  large  output  from  beds 
which  show  a  relatively  high  mining  cost  may  be  actually 
more  profitable  than  a  much  smaller  output  exclusively  from 
the  larger  and  cheaper  beds. 

Hence,  it  is  apparent  that,  even  from  the  standpoint  of 
immediate  profit,  it  may  be  advisable  to  mine  the  thinner  beds 
with  the  thicker,  and  considering  the  ultimate  yield  of  a  prop- 
erty, there  can  be  no  question  as  to  the  advisability  of  such 
a  course.  The  actual  minimum  minable  thickness  being 
dependent  upon  so  many  conditions  is  not  susceptible  to  gen- 
eral determination  and  should  be  studied  for  each  individual 
case. 

Conveyor  system. — The  method  of  operation  by  conveyors 
herein  described  has  been  in  use  in  a  number  of  collieries 
working  some  comparatively  thin  measures  in  one  of  the  coal 
fields  in  Scotland,  and  has  proved  its  success  and  applicability 
through  a  period  of  at  least  10  yrs.  In  some  respects  the 
method  adopted  was  unusual  in  that  while  conveyors  are  in 
operation  by  themselves  no  coal-cutters  are  employed,  under- 
cutting being  done  by  hand.  Present-day  practice  always  con- 
siders conveyor  work  an  adjunct  to  machine  mining;  but  here 
is  a  case  of  conveyor  practice  by  itself.  Another  distinctive 
feature  is  that  the  opening  and  development  of  the  mine  for 
additional  faces,  as  well  as  running  the  usual  longwall  faces 
with  the  conveyor,  are  being  done  with  a  conveyor  wall. 

The  coal  seam  on  the  average  is  3  ft.  9  in.  thick,  but  owing 
to  the  presence  of  stone  bands  is  rather  broken  up.  This 


94  COAL  MINING  COSTS 

means  that  after  removing  31  in.  of  coal  there  remains  about 
14  in.  of  stone  to  be  disposed  of.  The  thick  stone  parting 
near  the  bottom  of  the  coal  is  a  yellow-white  sandstone  that 
breaks  in  flat  squares,  eminently  suited  for  building  the  road 
pillars  in  longwall  working. 

In  working  under  the  old  system,  "docks,"  or  "deeps," 
were  driven  direct  to  the  dip,  the  inclination  being  8  deg. 
These  "dooks"  were  driven  in  the  solid  coal,  with  pillars 
turned  off  every  150  ft.  on  the  dip  and  60  ft.  on  the  level  course. 
Every  300  ft.  levels  were  broken  off  right  and  left,  and  a  long- 
wall  face  commenced  two  pillar  lengths  from  the  center  deep, 
in  a  fan-shaped  fashion,  which  as  it  opened  out  gradually 
edged  uphill  until  its  upper  corner  worked  along  the  waste 
of  the  level  above,  and  the  face  stretched  from  one  level  to 
another. 

In  driving  the  "deeps"  three  roads  were  allowed — one  for 
haulage  and  intake  air  current  while  the  two  on  either  side 
where  needed  for  return  air.  In  order  to  operate  the  long- 
wall  faces  at  low  cost,  slants  were  driven  uphill  from  the 
lowest  level,  and  from  the  slant  several  parallels  to  the  main 
bottom  level  turnecT  into  the  coal  face,  these  being  a  distance 
of  40  ft.  apart. 

During  the  ordinary  longwall  methods  of  working  the  fol- 
lowing men  were  employed  in  the  section:  Miners,  14; 
brushers,  7;  trackmen,  2;  timbermen,  3.  The  output  was  45 
tons. 

The  miners  trammed  their  own  coal  to  the  main  level.  In 
the  layout  of  the  workings  for  the  conveyor  there  was  little 
actual  difference  in  the  direction  of  development.  The  faces 
still  extended  across  the  strike;  but  in  place  of  the  three 
parallel  deeps  a  longwall  face  was  laid  out  at  an  angle  between 
the  dip  and  strike,  so  that  the  left-hand  end  was  the  most 
advanced,  which  allowed  the  coal  and  water  to  gravitate  to 
one  point.  The  driving  and  formation  of  pillars  were  thus 
dispensed  with,  and  the  longwall  system  of  extraction  became 
a  developing  system  as  well. 

The  conveyor  in  use  is  of  the  shaking  type,  and  has  been 
adapted  from  continental  practice.  The  height  from  the  floor 
to  the  edge  of  the  pan  is  only  9  in.  This  means  that  the  work- 
man is  required  to  raise  the  coal  only  slightly  over  this  height 


MINING  COSTS  95 

instead  of  a  former  29  in.,  which  accounts  for  more  work  with 
decreased  effort.  The  width  is  only  18  in.,  and  this  allows 
of  the  distance  from  coal  face  to  waste  line  of  props  being 
kept  at  a  minimum. 

The  principal  dimensions  of  the  driving  engine  are  as  fol- 
lows :  Horsepower  of  engine,  12 ;  air  consumption  at  60  strokes 
per  minute  and  60  Ib.  pressure,  18  cu.ft. ;  stroke  of  engine,  5  in. ; 
diameter  of  cylinder,  7  in. ;  weight  of  engine,  572  Ib. 

The  engine,  driven  by  compressed  air,  is  a  simple,  plain, 
air  cylinder,  with  broad  bed-plate,  which  may  be  bolted  to 
planks,  which  are  in  turn  wedged  against  the  roof  by  timber. 
The  total  width  is  18  in.,  and  the  length  24  in.  Connection 
to  the  conveyer  is  through  a  lever  action  and  rigid  attach- 
ment, the  cylinder  being  placed  at  right  angles  to  the  line  of 
the  conveyor. 

The  pack  walls  on  the  top  side  of  the  driving  road  are 
uniformly  built  in  line,  2  ft.  or  so  back  from  the  edge,  this 
space  being  utilized  for  the  engine.  Air  is  furnished  from  a 
power-driven  air  compressor  that  stands  in  a  small  recess  in 
the  side  of  the  road.  The  principal  dimensions  are  as  follows : 
Horsepower,  15 ;  r.p.m.,  960 ;  amperes,  16.5 ;  voltage,  500 ;  cycles, 
50;  air  pipe,  l1/^  in- ;  cylinders,  4;  strokes  per  minute,  air 
cylinders,  480;  pressure,  70  Ib. ;  space  occupied,  5  ft.  6  in.  by 
3  ft.  5  in.  The  air  is  passed  through  an  18-ft.  hose  to  the  air 
cylinder  of  the  conveyor. 

The  rate  of  advance  changed  from  160  ft.  in  six  months 
under  the  old  system  to  270  ft.  over  the  same  period  under 
the  conveyor  system.  This  is  not  remarkable  compared  to 
machine  working  advances,  but  it  represents  a  considerably 
increased  rate  of  extraction  for  handwork.  The  operation  of 
each  face  in  the  colliery  is  regular  and  at  the  same  rate,  the 
only  determining  factor  being  the  number  of  men  employed. 

Shifting  of  the  conveyor  takes  place  every  second  night, 
so  as  to  get  under  the  fresh  rock  as  soon  as  possible ;  but  this 
is  governed  by  the  rate  of  cutting  and  loading.  The  air  engine 
is  moved  at  the  same  time  as  the  conveyor,  but  the  compressor 
only  when  the  length  of  hose  is  reached.  The  operation  of 
shifting  is  accomplished  by  a  night  force  of  eight,  who  also 
shift  the  compressor  when  necessary  and  set  all  timber 
required. 


96 


COAL  MINING  COSTS 


A  comparison  of  the  costs  of  operation  and  performance 
accomplished  by  the  old  and  the  new  systems  has  worked 
out  much  as  follows: 


Hand 
Operation 

Conveyor 
Operation 

Tons  per  man 

3  2 

4  6 

Length  of  face  

300  ft 

300  ft 

Tons  per  section 

45 

55 

Length  of  face,  per  man  
Number  of  miners  .  . 

43  ft. 
14 

40-50  ft. 
12 

Number  of  deadwork  men  

12 

2 

Number  of  conveyor  men.  .  . 

1 

Shifting  conveyor  

.  .  .  (Average 

4 

Number  of  roads  to  maintain  

per  night) 

7 

2 

Tons  per  road  

Time  stripping  
Distance  bottom  level  in  advance  .  . 

/  6  .  4  Main  level 
I        Bottom  '  ' 
9      hr. 

49.00 
6.00 
9hr. 
40ft. 

COSTS 


Cutting 
Shooting    /                 

0  72 

0  72 

Loading    j 
Brushing                 

0  41 

(Done  by  squad 

Maintaining  roads  

0  08 

shifting  conveyor) 
0  08 

Tramming 

(Included    with 

(Included  for  low- 

Shifting conveyor  

shooting  and 
loading) 

er  level  in  ton- 
nage rate) 
0  15 

Operating  conveyor.         

0  03 

$1.21 

$0.98 

It  will  be  seen  that  the  saving  in  cost  finally  effected  is 
entirely  due  to  the  elimination  of  roof  troubles,  which  the  con- 
veyor system  made  possible.  There  are  now  installed  10  con- 
veyor faces  at  this  operation. 

Mining  machinery. — Man-power  is  about  the  most  expensive 
energy  purchasable.  We  pay  a  laborer,  say  $2  for  9  hr.  work. 
This  man  is  capable  of  exerting  continuously  about  one-eighth 


MINING  COSTS  97 

of  a  horsepower.  In  other  words,  we  have  secured  1%  hp.-hr. 
for  $2,  or  we  pay  about  $1.78  for  man-power  per  horsepower- 
hour. 

In  marked  contrast  to  this  high  cost  of  energy  is  the  cost 
of  current  delivered  to  the  motor  of  a  mining  machine  which 
should  not  exceed  2  to  2%c.  per  horsepower-hour. 

It  is  the  realization  of  this  discrepancy  between  the  cost 
of  power  developed  by  man  and  that  developed  by  a  steam 
engine,  for  instance,  that  is  driving  the  coal  industry  to 
employ  machinery  wherever  such  employment  is  possible. 
Furthermore,  it  is  frequently  the  case — as  in  undercutting, 
for  example — that  the  machine  does  the  work  better;  that  is, 
it  cuts  deeper  and  affords  less  resultant  fine  coal  than  when 
mining  is  done  by  hand. 

It  is  probable  that  most  operations  that  may  be  performed 
by  machinery  require  a  greater  expenditure  of  power  than 
would  the  same  operations  performed  by  hand;  nevertheless 
it  has  become  almost  axiomatic  that  it  is  economical  to  supplant 
manual  power  by  machinery  wherever  possible.  Consequently 
inventors  are  continually  striving  to  perfect  mining  machines, 
and  other  power-driven  devices  that  will  do  away  with  the 
employment  of  muscular  energy. 

Cutting  machines. — Mining  machines  now  produce  about 
65  per  cent  of  the  nation's  coal  output  as  compared  with  35 
per  cent  in  1907.  The  economies  over  hand  mining  may  be 
summed  up  as  follows:  First,  the  actual  cost  of  mining  is 
lower,  due  to  the  fact  that  the  greater  cutting  capacity  of 
the  machine  makes  possible  a  greater  output  with  a  given 
labor  cost;  second,  the  quality  of  the  product  is  superior,  due 
to  the  deeper  and  more  uniform  undercut  of  the  machine, 
which  increases  the  percentage  of  lump  coal  10  to  30  per  cent 
over  hand  mining  methods;  third,  the  mine  may  be  more 
rapidly  developed  due  to  the  much  greater  speed  with  which 
entries  can  be  driven  with  machines  insuring  a  rapid  return 
on  the  capital  invested;  fourth,  the  ability  to  mine  seams  in 
which  the  height  of  the  coal,  or  the  character  of  the  roof,  has 
prevented  mining  by  hand,  on  a  commercial  basis. 

During  the  period  from  1891  to  1914,  the  average  tons  of 
coal  mined  per  mining  machine,  per  year,  in  the  United  States 
was  about  13,700.  In  1913  there  were  2208  shortwall  mining 


98  COAL  MINING  COSTS 

machines  in  use  and  in  1914  there  were  3024,  an  increase  of 
32  per  cent  in  one  year.  In  1914  there  were  6859  breast 
machines  in  use,  which  is  an  increase  of  100  per  cent  over  the 
year  1904. 

One  of  the  most  noticeable  increases  in  coal  production 
mined  by  machines  has  been  in  Kentucky,  where,  in  1912, 
66.4  per  cent  of  the  coal  was  mined  by  machines,  while  in  1914 
the  production  mined  by  this  method  was  77.2  per  cent. 

In  the  accompanying  table  the  unit  of  efficiency  is  given 
for  the  total  production  of  bituminous  coal  in  the  United 
States  for  the  years  1891  to  1915.  There  is  also  a  column  show- 
ing what  percentage  of  the  coal  was  machine-mined.  By 
machine-mined  coal  is  meant  all  coal  won  by  the  use  of  any  of 


£ 


FIG.  47. — Curves  showing  per  capita  production  and  per  cent  of 
machine-mined  coal. 

the  following  types  of  machines:  Punchers,  radially  mounted 
punchers  and  chain  breast,  shortwall  and  longwall  cutters. 
The  figures  given  in  the  table  were  taken  from  the  reports  of 
the  Bureau  of  Mines  on  the  yearly  production  of  coal. 

From  these  statistics  it  is  at  once  apparent  that  the  increase 
in  the  number  of  tons  mined  per  day  per  man  has  corresponded 
closely  with  the  increase  in  the  percentage  of  machine-mined 
coal.  In  order  to  show  this  more  clearly  the  two  curves  shown 
in  Fig.  47  have  been  drawn.  The  upper  curve  shows  the  tons 
per  day  per  man  and  the  lower  the  percentage  of  machine- 
mined  coal.  From  these  two  curves,  unless  some  radical 
changes  are  made,  it  can  be  estimated  that  about  the  year 
1928  all  coal  will  be  machine-mined  and  that  the  efficiency  of 
the  miners  will  have  increased  to  about  4.9  tons  per  man  per 
day. 


MINING  COSTS 


99 


PROPORTION  OF  MACHINE  MINED  COAL  IN  THE  UNITED  STATES  AND  PRO- 
DUCTION PER  MAN 


Year 

Average 
No.  of 
Men 

No. 
Days 
Worked 

Total 
Tonnage 

Tons 
per  Man 
per  Day 

Per  Cent 
Machine- 
Mined 

1891 

205,803 

223 

117,901,238 

2.57 

5.26 

1896 

244,171 

192 

137,640,276 

2.94 

11.86 

1897 

247,817 

196 

147,617,519 

3.04 

15.35 

1898 

255,717 

211 

166,593,623 

3.09 

19.46 

1899 

271,027 

234 

193,323,187 

3.05 

22.74 

1900 

304,375 

234 

212,316,112 

2.98 

24.86 

1901 

340,235 

225 

225,829,149 

2.94 

25.61 

1902 

370,056 

230 

260,216,844 

3.06 

26.75 

1903 

415,777 

225 

282,749,348 

3.02 

27.58 

1904 

437,832 

202 

278,659,689 

3.15 

28.21 

1905 

460,629 

211 

315,062,785 

3.24 

32.82 

1906 

478,425 

213 

342,874,867 

3.36 

34.66 

1907 

513,258 

234 

394,759,112 

3.29 

35.11 

1908 

516,264 

193 

332,573,944 

3.34 

37.04 

1909 



379,744,257 



37.52 

1910 

555,533 

217 

417,111,142 

3.46 

41.72 

1911 

549,779 

211 

405,907,059 

3.50 

43.89 

1912 

548,632 

223 

450,104,982 

3.68 

46.80 

1913 

571,882 

232 

478,435,297 

3.61 

50.07 

1914 

583,506 

195 

422,703,970 

3.71 

51.70 

1915 

557,456 

203 

442,624,426 

3.91 

55.00 

Mining  machines  and  the  necessary  equipment  for  success- 
fully operating  them  at  the  average  colliery  cost  a  large  amount 
of  money.  If  this  is  injudiciously  spent  in  equipping  a  plant 
for  cutting  coal  with  machines,  overhead  charges  for  interest, 
maintenance,  depreciation  and  taxes  will  be  correspondingly 
heavy.  Conditions  may  justify  the  expenditure  in  order  to 
properly  recover  certain  coals  and  at  the  same  time  safeguard 
life  and  property;  but  such  moneys  should  be  carefully  and 
wisely  expended  and  then  only  after  exhaustive  analysis  of 
conditions  surrounding  the  proposed  operation.  The  fact  that 
one's  neighbor  mines  his  coal  with  machines  is  not  sufficient 
reason  for  one  to  so  equip  his  own  property.  Usually  each 
mine,  and  especially  each  coal,  has  its  peculiarities  that  deserve 
careful  consideration.  Many  pointers  and  suggestions  that  are 
worthy  of  serious  deliberations  may  be  had  from  the  man  at 


100  COAL  MINING  COSTS 

the  face.  Such  suggestions  only  cost  when  ignored  or  neg- 
lected. 

It  is  fair  to  assume  that  all  up-to-date  companies  maintain 
accounting  systems  that  are  a  criterion  by  which  they  may 
determine  approximately  the  relative  costs  of  pick-mined  and 
machine-mined  coal.  However,  there  are  many  angles  that 
afford  viewpoints  not  generally  considered  in  this  connection. 

Interest  on  investment,  maintenance,  depreciation  and 
taxes  on  all  extra  equipment  over  that  necessary  for  the  suc- 
cessful operation  of  the  property  as  a  pick  mine  should  properly 
be  charged  to  machine-mined  coal.  In  this  list  should  be 
included  all  extra  housing,  boilers,  boiler  settings,  boiler  fittings 
and  accessories  such  as  feed  pumps,  steam  headers,  etc.,  in 
addition  to  generating  units,  settings,  switchboards  and  acces- 
sories, transmission  lines  and  machines.  Further,  also,  under 
the  head  of  supplies  should  be  included  and  charged  to 
machine-mined  coal  all  extra  repairs,  fuel,  water,  oil,  tools 
and  office  supplies  over  and  above  that  necessary  for  the  suc- 
cessful operation  of  the  property  as  a  pick  mine;  and  under 
the  head  of  labor,  should  be  included  and  charged  to  machine- 
mined  coal  all  extra  expenditures  for  electricians,  wiremen, 
firemen,  oilers,  drivers,  tracklayers,  bit  sharpeners  or  black- 
smiths and  clerical  force  over  and  above  that  necessary  for 
the  successful  operation  of  the  property  as  a  pick  mine. 

Men  are  quite  frequently  required  to  timber  after  machines, 
clean  slate  and  refuse  at  switches  and  turns  on  account  of  the 
additional  space  required  for  machines  to  turn;  extra  drivers 
are  also  frequently  necessary  Hue  to  the  fact  that  they  are 
required  to  get  sharp  bits  to  the  machines  and  dull  bits  to 
the  shop,  must  occasionally  await  the  moving  of  a  machine 
thereby  losing  time  and  in  some  mines  they  must  drive  further 
for  their  loads  or  past  one  extra  place  out  of  every  three,  due 
to  the  fact  that  three  working  places  are  allowed  each  two 
loaders.  Extra  track  layers  are  frequently  required  for  the 
same  reason,  viz.,  that  they  have  more  track  to  keep  up  and 
over  a  larger  territory  due  to  the  fact  that  three  places  are 
allowed  two  loaders.  Bit  sharpener  or  extra  blacksmith  should 
properly  be  charged  to  machine-mined  coal  where  the  machine 
men  are  not  charged  for  smithing. 

Purchasing  agents  and  bookkeepers  spend  considerable  time 


MINING  COSTS 


in  ordering  machine  supplies,  checking  freight  bills  and  keep- 
ing track  of  supplies.  Delays  and  shutdowns  due  to  trouble 
with  boilers  or  generating  units  should  properly  be  charged 
to  machine-mined  coal  where  machines  are  responsible  for  such 
trouble.  As  an  example,  there  are  sometimes  delays  in  both 
hoisting  and  haulage  due  to  the  generating  plants  being  over- 
loaded by  reason  of  having  been  required  to  furnish  power  to 
the  machines. 

It  is  quite  possible  with  the  advice  of  the  machine  makers 
to  buy  equipment  that  will  suit  the  underground  conditions 
at  the  coal  face,  but  the  organizing  of  all  the  factors  that  make 
for  successful  operation  of  a  mine  to  the  new  conditions  created 
by  the  advent  of  the  machine  is  a  subject  conveniently  forgot- 
ten by  the  seller  of  the  apparatus,  and  often  not  considered 
by  the  operator.  To  buy  equipment  without  looking  into  this 
question  is  much  like  purchasing  an  automobile  for  which 
gasoline  cannot  be  readily  procured.  Consequently,  organiza- 
tion and  reorganization  of  the  mine  are  the  most  important 
factors  in  success.  Consideration  of  the  following  table  will 
serve  to  more  clearly  illustrate  this  point. 

CONTRAST  OF  HAND-  AND  MACHINE-MINING  CONDITIONS 


Under  Hand 
Conditions 

Under  Machine 
Conditions 

Tons  

100 

360 

Number  of  roads  to  be  kept  open 

18 

18 

Tons  per  road  

6 

20 

Men  in  section  ...                       

40 

59 

Timber  to  handle,  single  pieces 

3000 

12  108 

Cars  of  coal  

100 

360 

Rate  of  advance,  inches                   .    .    . 

12 

42 

It  is  evident  at  once  that  the  greatest  feature  is  the  traffic 
increase.  Instead  of  200  cars  a  day  to  deal  with  there  are 
now  720,  besides  an  additional  number  at  night.  Instead  of 
3000  pieces  of  timber  to  take  in  there  are  now  12,000  to  supply. 
Other  supplies  have  increased  due  to  the  use  of  the  machine, 
and  with  the  increase  in  traffic,  ventilation  of  the  mine  work- 
ings has  to  be  maintained  at  a  higher  state  of  efficiency. 

The  finest  machines  may  be  worthless  if  the  system  of  back- 


—  -?  "\ 

-1    V    ?    a      *> 

o 


COAL  MINING  COSTS 

ing  them  up  fails.  Ordinarily  in  all  coal  sections  the  amount 
of  work  done  is  limited  or  controlled  by  one  single  factor. 
This  may  be  the  amount  of  cars  provided,  the  size  of  the  sec- 
tion or  the  capacity  of  the  outside  haul.  In  machine  mining 
only  one  thing  should  control  the  section,  and  that  is  the  ton- 
nage produced  at  the  face  each  night.  Everything  should  be 
subordinated  and  coordinated  thereto. 

Before  the  men  leave  their  places  at  the  face,  the  coal  should 
all  be  squared  up  properly  so  that  the  machine  can  get  to 
work  promptly.  If  there  is  any  coal  left  behind  not  cleared 
up,  broken  down  but  not  loaded,  or  ' '  noses "  overhanging,  a 
man  should  be  sent  round  in  advance  of  the  machine  to  see 
that  all  these  obstructions  are  cleared  away.  This  extra  cost 
is  more  than  offset  by  the  gain  made  in  the  time  of  the  machine, 
as  well  as  by  the  elimination  of  the  risk  of  the  possible  loss 
next  day. 

The  tracks  by  which  the  machine  travels  from  place  to 
place  should  be  so  arranged  that  the  time  lost  in  travel  is 
reduced  to  a  minimum.  In  a  thick  coal,  where  the  machine 
cuts  a  relatively  high  tonnage  in  each  place,  there  is  more  coal 
to  set  against  this  waste  but  with  thin  coal  this  lost  time  runs 
up  alarmingly,  because  the  amount  of  coal  in  each  place  is 
small. 

It  is  obvious  that  all  machines  are  doing  their  best  work 
where  the  going  is  continuous.  Idle  time  and  idle  men,  or 
those  employed  on  unproductive  work,  mean  a  loss.  The 
traveling  from  place  to  place  results  in  the  loss  of  a  certain 
amount  of  productive  cutting  time,  and  it  should  therefore 
be  cut  to  the  minimum.  Tracks  and  curves  should  be  easy 
and  well  placed,  so  as  to  facilitate  traveling,  and  trouble  in 
this  connection  should  never  occur  twice  running  at  the  same 
spot.  Nightmen  should  be  at  work  on  the  bad  piece  of  track 
that  same  evening. 

Machine  supplies  should  be  kept  handy  to  the  face.  If  the 
section  is  a  long  one,  supplies  should  be  kept  at  several  points. 
Tool  chests  with  proper  keys  should  be  provided,  otherwise  the 
oil  is  often  found  to  have  disappeared,  together  with  necessary 
hammers,  keys  and  similar  tools.  Machinemen  should  be 
capable  of  making  reasonable  repairs  themselves,  and  the  mate- 
rials for  doing  so  should  be  kept  on  the  ground. 


MINING  COSTS  10$ 

There  should  be  no  hard  and  fast  rule  that  only  electricians 
are  to  repair  machines,  neither  should  every  handy  man  be 
allowed  to  try  his  hand  at  machine  troubles.  It  is  a  good  thing 
to  have  spare  machines.  Each  of  the  machines  underground 
should  be  taken  to  the  surface  at  regular  intervals  for  over- 
hauling. Hardly  any  class  of  machinery,  unless  it  is  the  pumps, 
receives  worse  usage  than  does  the  coal  cutter,  and  in  the  dark 
and  poor  light  underground  defects  may  develop  that  will 
never  be  noticed  until  it  is  too  late  to  make  the  proper  repair. 
All  machines  should  be  operated  steadily  on  the  surface  for  a 
number  of  hours  and  thoroughly  oiled  and  tuned  up  before 
being  returned  within  the  mine. 

Machinemen  should  be  at  their  machines  an  hour  or  so 
before  the  time  of  starting  work,  in  order  that  each  cutter 
may  be  overhauled  and  oiled,  the  bits  changed,  and  all  such 
details  given  the  proper  attention.  Spare  bits  should  always 
be  on  hand  and  any  that  have  been  removed  should  be  sharp- 
ened in  the  morning,  to  be  sent  down  again  at  night.  The 
proper  place  for  all  bits  except  those  in  stock  in  the  warehouse 
is  in  the  mine  and  not  in  the  blacksmith  shop.  One  smith 
should  sharpen  all  these  tools  the  first  thing  in  the  morning, 
regardless  of  any  other  work.  It  should  never  be  necessary 
to  stop  the  machine  to  hunt  for  cutters  or  telephone  to  the 
surface  at  two  in  the  morning. 

When  electric  current  is  used,  the  ratio  of  power  given  out 
by  the  cutter  motor  to  the  power  required  to  drive  the  gen- 
erator may  be  taken  roughly  as  70  per  cent,  while  with  com- 
pressed air  the  ratio  of  power  given  out  by  the  machine  to 
power  required  to  drive  the  compressor  may  be  taken  as  in  the 
neighborhood  of  35  per  cent. 

In  other  words  the  steam  consumption  of  a  compressed-air 
plant  for  coal-cutting  is  about  double  that  of  an  electric  plant. 
The  figure  70  per  cent  taken  for  the  electric  cutter  will  not 
differ  much  whether  the  installation  is  well  or  badly  designed, 
but  in  a  poorly  planned  and  badly  maintained  compressed-air 
plant,  a  considerably  lower  efficiency  will  be  obtained  than  the 
35  per  cent  taken  as  representative  of  a  moderately  good  instal- 
lation. 

Installation  and  operating  costs. — The  cost  of  a  5-machine 
plant  may  be  summarized  as  follows,  figures  as  of  1905 : 


204  COAL  MINING  COSTS 

Power  plant,  including  boilers,  air  compressor,  air 

receiver,  feed  pump  and  feed-water  heater $  5,181 .00 

Mining  machines,  including  all  accessories,  and 
freight 4,125.00 

Installation  of  power  plant,  including  freight,  com- 
pressor foundation,  boiler  settings,  wooden  boiler 
and  engine  house,  water  tank,  fittings,  piping, 
labor,  etc 2,260.00 

Pipe  lines  above  ground  and  in  the  mine,  with  all 

fittings,  labor  and  freight 2,484 . 00 


Total  cost $14,050. 00 

The  expense  of  maintaining  and  operating  this  plant  may 
be  approximately  estimated  as  shown  below  figures  as  of  1905 : 

Interest  (6  per  cent)  on  investment,  depreciation 
(10  per  cent),  repairs  and  renewals  on  machines 
and  power  plant  and  extensions  of  pipe  lines ....  $  2,250 . 00 

Fuel,  6£  tons  per  day  of  one  8-hr,  shift,  at  50c.  per 
ton,  and  oil  and  waste,  50c.  per  day;  per  year  of 
200  working  days 750.00 

One  engineer  at  $75  per  month;  one  machine  boss 
and  pipeman  at  $75;  one  blacksmith  to  sharpen 
picks  at  $60 2,520.00 


Total  maintenance $  5,520.00 

The  above  figures  are  considered  the  maximum,  so  that  in 
actual  practice,  the  cost  of  maintenance  of  the  plant  will  prob- 
ably be  lower. 

If  we  assume  the  pick  rate  at  this  mine  to  be  60c.  per  ton, 
and  the  differential  rate  for  the  machine  runner  to  be  one- 
fifth  of  the  hand  rate,  then  the  machine  rate  will  be  12c.,  and 
the  loading  rate  30c.  Adding  3c.  for  blasting  makes  45c.,  leav- 
ing a  margin  of  15c.  for  profit  and  payment  on  the  plant. 
With  an  output  of  700  tons  per  day  and  considering  200  days 
in  the  year  the  annual  gross  saving  will  be  $21,000.  If  we  now 
deduct  the  total  cost  of  plant,  including  maintenance,  from  this 
saving,  we  have,  $21,000  minus  $19,570  which  leaves  a  net 
profit  of  $1430  at  the  end  of  the  first  year. 

Cost  of  machine  mining. — It  is  practically  impossible  to 
compare  the  actual  cost  of  mining  with  the  various  types  of 
cutting  machines.  The  machine  that  would  show  a  consider- 


MINING  COSTS 


105 


able  saving  at  one  mine  might  prove  inefficient  at  some  opera- 
tion. However,  when  it  comes  to  comparing  machine  mining 
with  hand  mining,  there  is  no  difficulty.  The  most  important 
point  to  consider  is  the  size  of  the  differential  favoring  machine 
mining  over  hand  mining.  In  many  districts  this  differential, 
or  margin,  amounts  to  about  15c.  (in  1910)  and  it  is  out  of 
this  differential  that  the  operator  makes  his  profit  and  pays 
for  the  plant.  At  a  mine  producing  1000  tons  per  day  and 
having  a  15c.  margin  in  favor  of  machine  mining,  the  gross 
saving  would  be  $150  a  day,  or  $30,000  per  year  of  200  days. 
In  such  a  case,  the  company  can  maintain  its  output  with  20 
per  cent  fewer  men  than  are  required  when  hand  mining  is 
employed. 

Herewith  is  a  statement  showing  the  cost  of  machine  min- 
ing with  longwall  machines  at  an  operation  where,  owing  to 
the  tough  and  " woody"  nature  of  the  coal,  which  necessitated 
paying  the  miners  excessive  "allowances,"  the  average  cost  of 
mining  with  picks  was  between  60c,  and  63c.  per  gross  ton, 
figures  as  of  1910: 

COAL  CUT  BY  ELECTRICITY  AND  LOADED  BY  DAY  LABORERS 


Hours 

Rate 

Amount 

Cost  per 
Ton 

Cutting         

530 
531 
255 
260 
222 
2716 
3067 
270 

30c. 
20c. 
25c. 
20c. 
25c. 
20c. 
20c. 
30c. 

$  159.00 
106.20 
63.75 
52.00 
55.50 
543  .  20 
613  .  40 
81.00 

$0.0393 
0.0262 
0.0158 
0.0128 
0.0137 
0.1343 
0.1516 
0.0200 

Scraping 

Trackmen  

Trackmen         

Shooting 

Slate  

Loading       

Foremen  

Total  labor  

$1674.05 

$136.01 
36.00 
17.00 
5.00 
24.04 

$0.4137 

$0.0336 
0.0090 
0.0042 
0.0012 
0.0060 

SuDDlies 

Depreciation  on  machines          

Interest  6  per  cent  on  $3400 

Repairs  and  maintenance  (est 
Power 

[mated) 

4045  tons,  19  cwt.,  at  total 

cost  of 

$1892.10 

$0.4677 

106  COAL  MINING  COSTS 

The  above  cannot  be  taken  as  a  typical  or  average  state- 
ment, as  the  conditions  at  the  operation  referred  to  are  much 
less  favorable  to  pick  mining  than  at  most  other  operations 
in  this  field.  The  pick-mining  rate  along  New  River  is  50c. 
per  gross  ton,  and  at  most  of  the  operations  on  Piney  creek 
and  Loup  creek  it  is  40c.  per  gross  ton,  as  compared  with  69c. 
in  central  Pennsylvania,  so  that  the  economy  by  the  use  of 
machines  is  much  less  than  would  appear  from  the  figures 
above  quoted. 

The  following  are  detailed  figures  of  the  mining  cost  at 
two  operations,  one  where  all  of  the  coal  is  cut  by  electric 
chain  undercutters,  and  the  other  where  all  of  the  coal  is  cut 
by  air-driven  punching  machines.  The  pick-mining  rate  for 
the  year  when  these  statements  were  compiled  was  62c.  per 
gross  ton. 

ELECTRIC  BREAST  MACHINES,  OUTPUT  FOR  ONE  YEAR  321,808  GROSS  TONS 

Per  Ton 

Labor ; $0.4134 

Material 0.0259 

Insurance  and  taxes 0 . 0009 

Depreciation 0.0091 

Interest  charges 0.0025 


$0.4518 
COMPRESSED-AIR  PUNCHING  MACHINES,  OUTPUT  FOR  YEAR  99,207  GROSS  TONS 

Per  Ton 

Labor $0.4810 

Material 0.0211 

Insurance  and  taxes 0 . 0095 

Depreciation 0. 0132 

Interest  charges 0 . 0036 


$0.5284 

These  two  examples  cannot  be  taken  as  a  general  average, 
because  in  the  instance  quoted  where  chain  machines  are  used, 
the  mining  conditions  are  rather  exceptionally  favorable.  In 
the  other  case,  where  punching  machines  are  used,  the  con- 
ditions are  about  the  same  as  the  general  average  in  that  field. 

The  tonnage  produced  per  machine  of  a  given  feed  varies 
according  to  the  thickness  of  the  coal,  its  hardness,  the  width 
of  the  working  places  and  the  length  of  the  transfers. 


MINING  COSTS  107 

The  West  Kentucky  Coal  Co.  operating  nine  mines  in  the 
western  part  of  that  state  with  seams  ranging  from  4  ft.  7  in. 
to  8  ft.  6  in.  thick  gets  about  150  tons  per  machine-shift.  This 
is  an  average  of  nine  mines  and  includes  machines  used  on 
development  work  in  areas  practically  worked  out  where  the 
output  of  the  machines  is  naturally  limited.  In  one  seam  rang- 
ing from  5  to  8y2  ft.  thick  the  production  is  200  tons  per 
machine-shift.  The  record  cut  for  all  the  western  Kentucky 
field  up  to  1917  was  300  lineal  feet  of  face  in  9  hr. 

At  the  mines  of  the  W.  G.  Duncan  Coal  Co.  in  this  same 
field  where  five  machines  are  in  use  the  average  production  is 
nearly  250  tons  per  machine  shift.  At  these  mines  the  nature 
of  the  coal  is  such  that  they  are  able  to  use  a  25-in.  feed  over 
a  61/^-ft.  cutter  bar  or  a  21-in.  feed  over  a  7%-ft.  bar;  they 
are  also  getting  a  good  tonnage  per  bit  sharpened,  all  of  which 
factors  make  a  large  production  per  machine  possible. 

The  maintenance  cost  is  probably  the  most  important  item 
in  connection  with  the  mining  machine.  The  successful  and 
economical  operation  of  a  mining  machine,  like  any  other 
piece  of  machinery,  depends  largely  on  the  human  element. 
By  using  care  in  selecting  hostlers  who  will  later  become  run- 
ners, an  efficient  machine  organization  can  be  built  up.  There 
is  an  instance  where  one  man  operated  a  machine  continuously 
for  four  years  without  calling  on  the  machine  boss  except 
occasionally  for  some  small  repair  part  to  replace  one  actually 
worn  out. 

Unfortunately,  men  of  this  type  are  not  numerous,  and  for 
the  average  runner  some  incentive  is  necessary  to  sufficiently 
interest  him  in  getting  the  best  results  from  his  machine. 

At  the  mines  of  the  West  Kentucky  Coal  Co.  a  bonus  system 
is  in  effect  as  follows :  A  general  supply  stock  is  kept  at  each 
division  and  all  supplies  as  purchased  are  charged  to  this  sup- 
ply account.  One  machine  boss  has  charge  of  the  machines 
and  other  electrical  equipment  in  use  at  each  division,  with 
one  or  more  helpers  as  conditions  may  require.  All  supplies 
are  issued  by  either  the  machine  boss  or  his  helper  and  charged 
to  the  mining-machine  account  of  the  mine  where  used.  The 
average  maintenance  cost  for  the  year  1915  was  taken  as  a 
standard. 

The  maintenance  cost  per  ton  mined  for  the  year,  includ- 


108  COAL  MINING  COSTS 

ing  the  month  for  which  the  bonus  is  to  be  figured,  is  sub- 
tracted from  the  standard  described  above.  This  difference, 
multiplied  by  the  tonnage  of  the  month,  is  divided  equally 
between  the  company  and  the  machine  and  repairmen. 

It  can  readily  be  seen  that  in  this  way  the  men  can  make 
a  gradually  increasing  bonus,  and  this  they  have  succeeded 
in  doing.  At  the  same  time  it  is  impossible  by  skillful  manipu- 
lation of  supplies  on  the  part  of  machine  and  repairmen  to 
show  a  large  premium  in  one  month  followed  by  a  high  main- 
tenance cost  and  no  premium  the  succeeding  month,  which 
could  be  arranged  were  each  month  figured  separately. 

As  an  example,  to  show  how  this  premium  system  works 
out:  A  mine  having  a  machine  maintenance-cost  standard  of 
21/4c.  per  ton,  determined  as  above,  produces  in  a  given  month 
20,000  tons.  The  cost  per  ton  for  the  year,  including  the 
month  in  question,  is  1^2°^  making  a  net  saving  on  the  tonnage 
of  the  month  of  $150,  which  is  divided  as  follows :  $75  to  the 
company  and  $75  among  five  machine  runners,  five  hostlers  and 
one  repairman,  in  proportion  to  their  earnings  for  that  month. 
For  the  most  part  the  men  engaged  in  the  operation  of  the 
machines  have  taken  unusual  interest  in  reducing  maintenance 
costs,  and  the  men  at  the  different  mines  rival  one  another  in 
the  attempt  to  establish  the  best  record. 

The  cost  of  supplies  for  the  year  1915  was  0.96c.  per  ton. 
This  includes  bits,  bit  boxes,  cables  and  all  supplies  used  in 
the  operation  of  machines,  but  does  not  include  depreciation. 

In  this  connection  it  is  worth  notice  that  the  first  machine 
purchased  has  been  in  continuous  service  for  11  yr.  and  is  still 
in  as  good  operative  condition  as  one  just  out  of  the  factory. 

From  studying  the  results  of  several  machine  installations, 
particularly  those  of  the  United  States  Coal  &  Coke  Co.,  and 
the  Clinchfield  Coal  &  Land  Corporation,  it  was  found  that 
deep  undercutting  is  a  decided  advantage,  and  where  con- 
sistent with  the  mining  conditions,  nature  of  coal,  etc.,  should 
be  recommended  since  it  reduces  (1)  first  cost  of  machine 
installation,  (2)  powder  consumption,  (3)  cutting  cost  per  ton. 
Moreover,  where  a  certain  output  is  expected,  deep  undercut- 
ting should  reduce  the  territory  under  development.  This  is 
an  important  feature  when  the  maintenance  of  roads,  ventila- 
tion, and  supply  of  timber  are  taken  into  consideration.  Fur- 


MINING  COSTS 


109 


thermore,  it  should  increase  the  percentage  of  lump  and  give 
the  loader  a  more  definite  or  dependable  quantity  of  coal  down, 
thus  increasing  his  efficiency,  and  thereby  raising  that  of  the 
entire  plant. 

Comparative  cost  of  alternating  and  direct  current  for 
machines. — In  a  mine  where  alternating-current  machines  are 
to  be  utilized,  large  tonnage  should  be  developed  in  as  small 
an  area  as  possible  in  order  to  secure  the  most  economical  use 
of  the  machines.  This  is  shown  in  the  accompanying  diagram 
of  a  typical  room-and-pillar  working,  Fig.  48. 


High  Tension  Unes(Cable 

•  (3WRC.Wires) 


FIG.  48. — Typical  room-and-pillar  development  using  alternating- 
current  machines. 

In  this  mine  the  rooms  available  for  the  mining  machines 
are  possibly  far  in  excess  of  what  the  mining  machine  is  capable 
of  cutting.  In  territory  distributed  as  in  this  example  one 
mining  machine  should  cut  at  least  7  rooms  in  a  working  shift 
or  14  rooms  on  a  double  shift.  This  would  give  approximately 
300  lin.ft.  of  coal  per  shift,  depending,  of  course,  upon  cir- 
cumstances, the  above  example  being  under  average  working 
conditions. 


110  COAL  MINING  COSTS 

The  'power  cost  of  operating  a  single  alternating-current 
mining  machine  is  comparatively  low,  as  with  the  arrange- 
ment laid  out  in  the  diagram  one  mining  machine  would  not 
use  over  2000  to  3000  kw.-hr.  per  month  working  single  shift. 
This  small  consumption  of  power  is  principally  due  to  the 
mining  machine  having  sufficient  voltage  at  all  times,  and, 
therefore,  working  at  its  highest  efficiency. 

One  mining  machine  in  4-ft.  coal  and  with  a  7%-ft.  cutter- 
bar  mining  say,  300  tons  to  the  shift,  or  6000  tons  a  month, 
will  have  a  power  expenditure  of  less  than  y2  kw.-hr.  to  the 
ton.  This  would  be  less  than  Ic.  per  ton  on  the  average  rate 
of  central-station  contracts. 

A  comparison  of  maintenance  between  the  alternating-  and 
direct-current  machines  shows  favorably  for  the  former.  With 
direct-current  power  in  use,  especially  with  a  mine  that  is 
supplied  from  some  isolated  power  house  on  the  outside  of  the 
mine  and  a  considerable  distance  from  the  workings,  the 
voltage  is  usually  low;  consequently,  much  armature  trouble  is 
experienced  in  the  motor  on  the  cutting  machine. 

With  an  alternating-current  mining  machine  served  from 
central-station  power  there  is  an  assurance  of  good  voltage, 
as  at  no  place  need  the  machine  be  at  a  greater  distance  than 
1500  ft.  from  the  distributing  transformers.  This  distance  in 
any  mine  will  give  the  mining  machine  more  territory  than  it 
is  possible  for  it  to  work. 

Alternating-current  power  is  probably  more  economical  for 
general  mining  use,  other  than  haulage.  In  fields  where  cen- 
tral-station power  is  available  pumps,  hoists,  fans  and  tipple 
equipment  are  all  run  by  alternating-current  power  and  appear 
to  give  more  satisfactory  results  than  direct  current. 

In  mines  where  alternating  current  is  available  and  electric 
haulage  is  required  it  would  be  an  added  expense  to  install 
wire  for  alternating-current  cutting  machines,  as  the  con- 
ductors that  supply  power  for  haulage  can  be  used  for  direct- 
current  cutting  machines.  This  accounts  for  the  general  use 
of  direct-current  cutting-machines  in  the  large  mines  where 
electric  haulage  is  employed. 

To  the  small  operator  with  limited  capital  the  alternating- 
current  mining  machine  has  shown  the  way  to  a  greatly  in- 
creased production  without  the  large  investment  formerly 


MINING  COSTS  111 

required.  It  is  possible  to  decrease  mining  costs  and  increase 
production  at  a  comparatively  small  expenditure  above  the  cost 
of  the  machine  itself. 

Arc  wall  cutters. — The  Jeffrey-Drennen  adjustable-turret 
coal  cutter,  was  designed  to  make  a  cut  at  any  elevation  desired. 
This  was  installed  at  Jenkins,  Ky.,  where  the  coal  seam  varies 
from  6  to  8  ft.  in  thickness.  It  is  clear,  bright  and  free  from 
sulphur  or  other  impurities  with  the  exception  of  a  band  of 
shale  located  at  a  height  of  from  2  to  5  ft.  from  the  bottom. 
This  varies  in  thickness  from  nothing  to  19  in. 

With  the  customary  methods  of  undercutting  it  would  be 
impossible  when  shooting  to  prevent  this  shale  from  becoming 
mixed  with  the  coal,  but  by  the  use  of  a  machine  adapted  to 
cutting  out  or  removing  this  parting  before  the  coal  is  shot 
down  this  difficulty  is  overcome. 

The  machine  is  mounted  on  a  turnable  truck,  which  carries 
four  heavy  standards  or  uprights,  on  which  the  machine  proper 
is  raised  or  lowered  or  adjusted  to  the  desired  height  at  which 
to  cut  out  the  dirt  seam.  The  cutter  is  designed  for  a  minimum 
height  of  2  ft.  from  the  bottom,  and  can  be  adjusted  to  cut 
at  any  position  between  2  and  5  ft.  The  raising  or  lowering  of 
the  machine  is  accomplished  by  power  through  a  disk  friction 
clutch,  which  enables  the  operator  to  absolutely  control  the 
elevation  of  the  machine  to  a  nicety,  3  ft.  of  a  vertical  move- 
ment being  accomplished  in  about  25  seconds. 

The  cutting  is  done  in  the  shale  at  the  bottom  of  the  band 
with  the  lower  nose  of  the  bits  cutting  into  the  coal  about 
!/4  in.,  which  causes  the  shale  to  fall  down  in  the  kerf,  after 
which  it  is  cleaned  out,  loaded  in  cars  and  hauled  out  of  the 
mine.  This  insures  an  absolutely  clean  product. 

A  15-ft.  glace  can  be  cut  in  11  min.  from  the  time  the 
machine  enters  the  room  until  it  is  ready  to  leave.  Twenty- 
five  rooms  have  been  cut  in  a  shift  of  10  hr. 

The  Utah  Fuel  Co.  of  Somerset,  Colo.,  cut  258  lin.  ft.  of 
coal  in  2  hr.  24  min.  with  this  machine.  The  vein  is  14  ft. 
thick. 

Post  punchers. — Compressed-air  post  mining  machines  of 
the  radial  type  were  installed  at  the  Pacific  Coast  Coal  Co.'s 
mines  about  1909,  and  it  was  found  that  after  a  mining  was  put 
in  with  these  machines  the  coal  could  be  sent  down  the  chutes 


112  COAL  MINING  COSTS 

with  only  a  little  pick  work,  and  without  any  powder,  except 
an  occasional  light  shot  at  a  corner  or  to  shoot  out  a  "  nigger 
head.'*  In  fact  the  powder  consumption  was  reduced  over 
95  per  cent. 

The  lump  coal  was  increased  in  this  way  from  about  25 
per  cent  to  about  60  per  cent;  and  the  practical  elimination  of 
powder  made  the  mine  much  safer. 

The  rooms  are  driven  45  to  50  ft.  wide.  The  first  cut  is 
made  from  a  post  set  7  ft.  from  the  left  rib  and  about  18  in. 
from  the  face.  A  cut  8  ft.  in  depth  is  put  in,  using  an  exten- 
sion bar  80  in.  long.  The  chuck  enters  the  cut,  which  accounts 
for  the  mining  being  deeper  than  the  length  of  the  extension. 
After  the  80-in.  extension  has  been  swung,  a  100-in.  bar  is 
used  to  square  up  the  cut.  One  man  operates  the  machine, 
swinging  it  by  means  of  the  worm-crank  with  one  hand,  and 
feeding  the  cylinder  forward  two  or  three  turns  with  the  other 
at  each  end  of  the  swing. 

The  machine  and  the  posts  remain  in  the  room  at  the  face 
until  the  room  is  completed,  for  there  is  no  shooting  of  the 
coal  that  can  injure  the  machine  nor  any  loading  (as  in  a 
flat  seam)  that  the  machine  would  interfere  with.  Hence 
there  is  no  waste  of  time  due  to  moving,  except  from  post  to 
post,  and  little  heavy  lifting.  Two  men  can  set  up  the  machine, 
with  ease,  as  the  heaviest  parts  (the  machine  and  shell)  weigh 
only  225  Ib. 

These  machines  will  average  about  300  sq.  ft.  in  an  3-hr, 
shift,  or  from  45  to  55  tons  per  machine  per  shift,  according 
to  the  height  of  the  coal.  While,  as  stated,  the  reason  for  the 
installation  of  these  machines  was  solely  to  increase  the  pro- 
portion of  lump  coal,  even  if  the  cost  of  mining  was  increased, 
the  results  point  strongly  toward  a  material  reduction  in  the 
cost  of  mining,  after  all  interest,  depreciation,  power,  pipe 
line,  and  maintenance  charges  have  been  made  against  the 
machines. 

In  shooting  off  the  solid,  the  former  method,  a  yardage 
system  of  payment  was  used,  the  rate  being  $9.50  for  a  50-ft. 
room.  Three  men  worked  together,  furnishing  their  own  pow- 
der. The  reason  for  using  a  yardage  and  not  a  tonnage  system 
was  because  of  the  impurities  in  the  seams,  and  the  pitch. 

Repair  costs. — It  pays  well  to  have  the  machinery  in  shape 


MINING  COSTS  113 

to  run  a  full  day  when  the  mine  is  working.  By  this  means 
whatever  men  are  in  the  mine  are  given  a  chance  to  produce 
some  coal,  and  the  cost  of  production  is  considerably  cheapened 
because  repair  costs  are  reduced  to  a  minimum.  This  is 
important,  for  the  cost  for  repairs  is  too  high  in  almost  every 
mine. 

Furthermore,  by  giving  each  machine  the  needed  care  there 
will  be  an  increase  in  tonnage  which  will  permit  of  a  further 
saving  in  the  cost  of  production.  In  order  to  lower  the  repair 
cost  and  increase  the  efficiency  of  the  machinery,  it  is  neces- 
sary to  have  a  system  simple  but  accurate,  which  will  give  an 
individual  record  of  the  performance  and  expense  of  each 
piece  of  machinery. 

There  should  be  a  book  kept  at  each  mine  by  the  electrician 
for  the  purpose  of  recording  the  working  hours  of  each  machine, 
and  in  this  the  hours  idle  should  be  marked  down.  With  a 
little  attention  it  will  be  possible  to  get  the  average  number 
of  tons  or  cars  each  machine  is  capable  of  producing,  and 
thus  it  will  be  possible  to  find  at  the  end  of  each  month  the 
number  of  tons  lost  through  the  inefficiency  of  any  machine. 

Every  night  the  machineman  should  fill  out  a  report  show- 
ing the  make  and  number  of  his  machine  and  the  hours  it  has 
been  delayed,  and  he  should  state  any  defects  he  may  have 
noticed  in  its  operation.  The  electrician  should  then  have  these 
defects  repaired  and  sign  the  report  and  forward  it  to  the 
chief  electrician.  The  mine  electrician  should  also  fill  out  a 
daily  report  on  this  order: 


Min 

e  

NAM: 

B  OP  COMPANY 
Dai 

Hours 
of 
Labor 
3 
2 

Wage 
Rate 
25 
40 

Make  and  Number 
of  Machine  or 
Locomotive 
SS  No.   1620  Mining 
machine 

Repair  Parts 
or  Material 
Used 
1  worm  gear 
lkeyiXiX6in. 

Cost 
of  Part 
$14.50 
0.06 

State  if  Broken  or 
Worn     Out     and 
Cause 
Broken  teeth  worn 
too  thin  to  stand 

strain 

REMARKS. — Advise  that  this  machine  be  changed  to  easier  work  as  the  section  is  full 
of  rolls  and  the  machine  is  too  light  for  that  work. — S.  G.  MILLER. 

This  report  should  be  sent  daily  to  the  chief  electrician, 
who  would  have  each  item  charged  against  the  machine  on 
which  it  was  used.  Thus  at  the  end  of  each  month  the  exact 
cost  of  labor  and  material  used  on  each  machine  could  be 
easily  found. 


114  COAL  MINING  COSTS 

The  wiremen  and  bondmen  should  report  the  section  in 
which  they  have  been  working,  the  kind  of  work  they  did  and 
the  material  used.  This  should  be  recorded,  and  it  could  then 
be  seen  at  a  glance  how  much  copper  or  other  material  was 
used  to  work  out  a  section. 

The  mine  electrician  should  at  the  end  of  each  month  send 
a  report  to  the  chief  electrician,  showing  the  horsepower, 
make  and  number  of  each  motor,  machine  or  pump  at  the  mine, 
the  number  of  hours  in  use,  the  amount  of  time  out  of  com- 
mission for  repairs  and  the  amount  of  oil  used  on  it.  This, 
with  the  electrician's  daily  report,  gives  the  chief  electrician 
a  chance  to  get  down  to  facts. 

For  example,  if  a  certain  mining  machine  gives  much 
trouble,  he  can  see  at  a  glance  just  what  part  was  at  fault 
each  time,  and  with  these  facts  on  hand  he  can  proceed  to 
investigate  and  remedy  the  trouble.  Again,  if  a  certain  part 
wears  out  on  each  machine  of  a  kind,  he  will  know  that  it  is 
a  weak  part  in  the  construction  of  that  machine,  and  he  can 
take  it  up  with  the  manufacturer  who  will  probably  be  able 
to  suggest  some  way  to  overcome  it. 

The  mine  electrician  should  order  his  supplies  once  a  month, 
and  they  should  be  charged  against  him  and  not  against  the 
repair  cost  until  they  are  actually  used  and  charged  on  his 
report.  At  the  end  of  each  month  he  should  take  an  inventory 
of  all  the  supplies  he  has  at  the  mine  and  should  be  credited 
with  them. 

For  example,  he  has  $500  worth  of  supplies  on  hand  on 
the  first  of  the  month,  and  his  requisition  shows  that  he  has 
received  other  supplies  worth  $500.  At  the  end  of  the  month, 
on  taking  his  inventory  he  finds  he  has  $400  worth  left.  This 
shows  he  has  used  $600  wortK  of  material.  On  checking  up  his 
daily  reports  they  should  balance  with  this  figure  to  show  that 
everything  had  been  charged  in  its  proper  place. 

The  mine  foreman  should  mark  the  delays  to  the  machinery 
on  his  report  also,  and  this  should  check  with  the  mine  elec- 
trician's and  efforts  should  be  made  to  keep  them  correct. 

Each  piece  of  machinery  should  be  treated  as  a  workman. 
The  time  in  use,  the  amount  lost,  tons  of  coal  handled  or  cut, 
cost  of  labor  and  supplies  should  all  be  completely  recorded. 
This  will  show  which  is  the  most  efficient  machinery  to  buy 


MINING  COSTS  115 

for  future  use  and  which  to  discard,  what  parts  wear  longest 
and  what  parts  are  most  liable  to  give  out  first.  It  also  shows 
the  amount  of  material,  such  as  wire  and  hangers,  bond,  etc., 
in  each  section  of  the  mine. 

In  order  to  get  results  from  this  system  it  is  necessary  to 
have  skilled  and  competent  mechanics.  One  cannot  reduce 
costs  if  the  men  fail  to  do  their  share.  It  is  too  frequently 
the  case  that  companies  using  first-class  machinery  have  in- 
experienced and  underpaid  mechanics  tending  their  expensive 
equipment,  while  other  companies  having  a  poor  grade  of 
machinery  have  good  mechanics. 

Many  coal  companies  will  buy  an  expensive  piece  of  machin- 
ery and  then  let  an  inexperienced  man  experiment  with  it. 
The  result  is  the  repair  cost  is  large  for  the  first  few  years, 
though  it  ought  to  be  practically  negligible.  There  is  nothing 
gained  by  letting  someone  experiment  with  machinery,  and 
it  is  a  costly  practice.  Through  ignorance,  most  of  such  high 
repair  costs  are  blamed  on  the  equipment,  whereas,  if  the 
machinery  had  been  used  in  the  right  place  and  treated  as 
it  should  have  been,  there  would  have  been  no  trouble. 

The  average  cost  of  repairs  per  ton  of  coal  mined  is  about 
lOc.  while,  with  careful  attention  on  the  part  of  the  chief 
electrician  and  the  mine  electricians,  it  would  be  an  easy  mat- 
ter to  bring  it  down  to  5c.  per  ton,  and  if  all  did  their  best, 
and  the  machines  are  in  first-class  condition  it  should  be  brought 
to  less  than  3c.  with  no  interruptions  to  service  during  work- 
ing hours. 

Machine  bits. — The  bit  question  is  an  important  problem 
in  connection  with  the  operation  of  machines,  for  the  bit  is 
to  the  mining  machine  what  the  tooth  is  to  the  crosscut  saw. 
When  using  chisel-point  bits  in  connection  with  up  and  down 
pick  points,  the  sharpening  of  these  three  classes  of  bits  and 
their  distribution  to  ftie  different  machines  is  quite  difficult. 
The  Sullivan  Machinery  Co.  recognized  this  difficulty  and  fol- 
lowing the  old  5-position  chain  it  later  introduced  the  9  posi- 
tion superdreadnought.  Straight  pick-point  bits  in  this  chain 
give  a  good  clean  kerf,  making  comparatively  coarse  machine 
cuttings,  while  placing  on  the  machine  only  a  moderate  load. 

Formerly  all  bits  were  sharpened  by  hand,  but  this  is  a  slow 
and  most  expensive  process.  Small  trip  hammers  will  do 


116  COAL  MINING  COSTS 

more  than  any  other  one  thing  to  ease  bit  troubles.  One  man, 
with  a  boy  to  heat  the  bits,  can  sharpen  2500  to  3000  bits  per 
day,  making  on  an  average  a  better  bit  than  a  man  will  make 
without  the  hammer,  because  a  blacksmith  sharpening  by  hand 
will  endeavor  to  get  the  bits  as  hot  as  possible  so  as  to  save 
hammering.  In  doing  this  he  not  only  makes  a  badly  tem- 
pered bit  but  burns  away  valuable  bit  material. 

The  die  used  in  these  hammers  may  be  a  home  product  that 
will  probably  be  equal  to  anything  on  the  market  and  that  can 
be  made  for  $20  per  set.  For  tempering  compounds  a  solution 
of  cheap  laundry  soap  and  soft  water  gives  good  results  at 
little  expense.  One  cake  of  soap  is  used  to  25  gal.  of  water, 
washing  powder  being  added  where  the  water  is  too  hard  to 
lather  freely. 

At  the  beginning  of  the  shift  this  mixture  is  heated  to  the 
boiling  point  and  the  sharpened  bits  while  at  a  cherry-red 
heat  are  thrown  into  the  tempering  tub,  the  hot  bits  keeping 
the  mixture  boiling,  insuring  a  slowly  cooled  and  evenly  tem- 
pered bit.  The  object  is  to  furnish  each  machine  with  plenty 
of  bits  and  thus  encourage  the  runner  to  change  them  before 
they  become  dull  enough  to  load  the  machine.  It  is  cheaper 
to  sharpen  bits  frequently  than  to  run  the  risk  of  burning  out 
armatures  and  wearing  out  cutter  chains. 

In  Nos.  9  and  11  seams  the  West  Kentucky  Coal  Co.  uses 
approximately  150  sharp  bits  to  cut  100  tons  of  coal  while  in 
the  No.  12  seam  the  tonnage  per  bit  is  more  than  double  this 
amount.  The  number  of  bits  required  may  seem  high  in  this 
instance  but  it  was  frequently  the  case  in  these  mines  that  150 
bits  would  be  dulled  in  a  single  room  where  rock  or  heavy 
sulphur  bands  were  encountered.  In  spite  of  these  conditions 
the  records  over  three  years  at  these  mines  show  an  average 
production  of  35,000  tons  per  1000  bits  purchased. 

Loading  machines. — Because  of  the  decrease  in  the  supply 
of  labor  throughout  the  country  and  the  increasing  wages 
more  interest  is  being  centered  on  ways  and  means  for  reduc- 
ing the  amount  of  muscular  energy  required  in  the  mining  and 
loading  of  coal  underground.  Undercutting  and  loading  are 
probably  the  two  jobs  least  sought  for  in  American  coal  mines, 
and  it  is  increasingly  difficult  to  procure  men  who  will  per- 
form this  kind  of  work, 


MINING  COSTS  117 

Many  machines  have  been  designed  for  the  loading  of  coal. 
Some  of  them  both  mine  and  load  the  material,  while  others 
merely  load  it.  In  general,  no  difficulty  has  been  encountered 
in  devising  an  apparatus  to  place  the  coal  upon  the  cars,  the 
main  disadvantage  of  such  devices  being  that  it  is  impossible 
to  feed  cars  to  the  machines  with  sufficient  rapidity  to  make 
mechanical  loading  practicable.  As  a  result  such  machines 
stand  idle  a  goodly  portion  of  the  working  day  waiting  for 
the  loaded  cars  to  be  taken  away  and  their  places  filled  by 
empties.  Furthermore,  these  machines  are,  as  a  rule,  cumber- 
some and  difficult  to  move  because  of  their  great  weight. 

In  hand  shoveling  the  height  through  which  the  shovel 
must  be  raised  determines  the  capacity  of  the  loader.  With 
the  same  expenditure  of  muscular  energy  to  raise  the  coal,  a 
man  will  load  29  tons  into  a  wagon  52  in.  above  the  rail,  44 
tons  into  a  wagon  32  in.  above  the  rail,  and  140  tons  into  a 
conveyor  8  in.  above  the  floor,  allowance  being  made  for  clear- 
ance in  each  case.  Also  it  must  be  remembered  that  in  raising 
the  shovel,  the  foot  pounds  expended  in  raising  the  body  is 
greater  than  the  foot  pounds  exerted  in  lifting  the  coal.  When 
loading  into  a  conveyor  in  connection  with  a  loading  machine, 
the  coal  does  not  have  to  be  lifted  more  than  a  few  inches; 
some  of  it  can  be  rolled  or  pushed  on,  and  there  is  an  important 
saving  in  muscular  energy  effected. 

The  Jeffrey  loader  weighs  one  ton.  It  has  the  motor  and 
machinery  located  in  the  center  of  the  conveyor  over  the  sup- 
porting rail,  and  therefore  one  man  is  able  to  bear  down  on 
the  back  end  and  slew  the  machine  around  at  will,  and  one 
man  can  readily  push  the  loader  along  the  supporting  rail  to 
any  desirable  position.  The  supporting  rail  is  ordinarily  placed 
on  two  wooden  horses.  The  loading  machine  is  made  strong 
enough  to  admit  of  shooting  the  coal  down  on  top  of  the  front 
end  of  the  conveyor. 

It  is  provided  with  a  self-propelling  truck  to  move  from 
one  place  to  another. 

For  best  results  with  the  loading  machine  the  mine  should 
be  laid  out  systematically  with  just  enough  loaders  in  each 
section  to  keep  one  under  cutting  machine  busy.  For  instance, 
in  Fig.  49  is  shown  a  system  of  ten  rooms.  In  rooms  1  to  5 
are  loading  machines;  in  rooms  6  to  10  a  shortwall  machine. 


118 


COAL  MINING  COSTS 


Each  room  is  loaded  out  in  half  a  shift,  therefore  all  ten  rooms 
are  undercut  and  loaded  out  once  a  day.  One  gathering  loco- 
motive will  handle  cars  for  the  five  loading  machines,  if  fairly 
good  size  cars  are  used. 

The  Jeffrey  pit  car  loader  is  a  simple  conveyor,  driven 
by  an  electric  motor,  so  located  as  to  add  to  the  stability  of 
the  machine. 

Its  novelty  would  appear  to  lie  in  its  cost,  which  is  about 
one-seventh  that  of  a  shoveling  machine.  It  will  doubtless 
appeal  to  those  who  contend  that  coal  mining  is  attended  with 
too  many  delays  to  tie  up  large  amounts  of  capital  in  expensive 
machinery,  and  attract  those  who  fear  that  the  mine  is  no 
place  for  a  machine  designed  to  pick  up  the  coal. 


FIG.  49. — System  of  mining  suggested  for  adoption  with  coal  loading 

machines. 

The  claim  of  its  maker  that  the  output  from  a  room  can 
be  doubled  by  the  use  of  this  machine  appears  to  be  conserva- 
tive. The  fact  that  it  requires  only  two  men  to  operate,  whereas 
other  machines  require  from  four  to  eight,  is  vastly  in  its 
favor,  for  when  the  inevitable  delays  occur,  it  is  not  difficult 
for  two  men  to  find  useful  work  in  posting,  extending  track, 
etc.  Quite  a  number  of  these  machines  are  in  operation  at  the 
mines  of  the  Dominion  Coal  Co.,  Canada,  some  of  them  for  a 
period  of  three  years  or  more,  and  it  has  been  proven  that  two 
men  will  average  from  50  to  75  tons  per  shift. 

Three  men  using  this  machine  can  load  out  the  coal  from  a 
16-ft.  entry,  6-ft.  undercut,  in  one  hour,  where  the  height  of 
coal  is  42  in.  with  a  2  in.  to  6  in.  parting,  the  dirt  from  which 
has  to  be  picked  out. 

Three  men  have  loaded  7  cars  of  slate  in  35  min.,  the  capacity 
of  car  being  3400  Ib.  of  coal. 


MINING  COSTS  119 

In  5  ft.  coal  with  4  in.  binder,  two  rooms  were  drilled,  shot 
and  loaded  out  in  8  hr.,  making  39  cars  of  coal,  each  holding 
3300  Ib.  About  6  in.  of  draw  slate  and  binder  were  cleaned 
out  and  gobbed.  The  cost  of  this  machine  in  December,  1921 
was  $1500. 

The  Westmoreland  loader  weighs  10  tons,  is  electrically 
operated  and  self-propelled.  The  truck  is  rigid,  with  12-in. 
wheels,  one  front  wheel  loose  axle,  4-ft.  wheel  base,  and  when 
in  operation  the  truck  is  clamped  rigidly  to  the  rail.  A  3-ft. 
geared  turntable  rests  on  the  truck  frame,  and  on  this  is 
mounted  a  solid  steel  case,  12  ft.  long,  32  in.  wide  and  I2y2  in. 
high,  which  swings  free  above  the  wheels  through  an  arc  of 
160  deg.,  80  deg.  right  and  80  deg.  left,  the  width  of  sweep 
being  16  ft. 

The  steel  case  contains  a  ram  in  the  form  of  a  chute,  2  ft. 
wide,  6  in.  deep  and  above  the  floor,  and  with  an  adjustable 
shovel  end  at  the  front.  The  ram-chute  hangs  on  rollers  and 
is  moved  forward  and  backward  by  a  rack  and  pinion  at  the 
top.  This  represents  the  unique  feature  of  the  machine.  The 
ready  horizontal  movement  of  the  ram,  together  with  its  sweep 
of  8  ft.  on  either  side  of  the  machine,  enables  it  to  keep  close 
to  the  loose  coal  throughout  the  full  16  ft.,  or  slightly  more, 
of  width  and  thus  facilitates  the  work  of  the  laborer. 

The  adjustable  shovel  end  of  the  machine  takes  care  of 
undulating  bottom.  It  is  rigid  in  operation  for  all  sizes  of 
coal,  fine  slack  up  to  200-lb.  lumps  being  picked  up  at  the  rate 
of  one  ton  per  minute  by  the  chain  conveyor  in  the  ram.  The 
actual  rated  capacity  of  the  loader  under  fair  conditions  is 
20  tons  per  hour.  Five  men  are  required  for  its  operation, 
including  the  driver.  The  machine  will  work  in  seams  that 
are  not  less  than  4%  ft.  thick,  and  will  pass  over  curves  of  as 
low  as  12-ft.  radius.  The  total  power  requirement  is  16  hp., 
consisting  of  three  reversible  motors. 

The  Evans  scraper  loader  is  an  adaptation  of  the  old  main- 
and-tail  rope  principle  to  the  operation  of  a  modified  scraper 
as  a  means  for  conveying  the  coal  from  the  face  and  out  to 
the  room  neck,  at  which  point  it  is  dumped  into  the  car. 

The  apparatus  consists  of  a  double  drum  hoist,  which  may 
be  driven  by  either  air  or  electric  power,  two  wire  ropes  of 
suitable  lengths,  and  a  V-shaped  bottomless  scoop,  or  drag.  ID 


120 


COAL  MINING  COSTS 


addition  to  this  major  equipment,  each  room  is  provided  with 
deflectors,  loading  pans  and  aprons.  All  coal  is  loaded  on  cars 
in  the  entry,  the  plan  being  to  set  a  trip  of  empty  cars  above 
the  room  neck  so  that  no  time  will  be  lost  in  spotting  another 
car  when  one  is  loaded. 

It  is  figured  that  with  rooms  of  300  ft.  length  it  is  possible 
to  load  at  least  nine  tons  of  coal  per  hour.  The  average  haul 
in  this  case  amounts  to  150  ft.,  and  the  hoist  being  geared  for 
a  rope  speed  of  300  ft.  per  minute  the  traveling  should  be  done 
in  one  minute  actual  running  time.  Assuming  that  a  minute 
is  lost  at  the  face  and  another  minute  at  the  loading  point, 
there  would  be  20  trips  per  hour,  which  on  a  basis  of  900  Ib. 
per  scoop  amounts  to  nine  tons. 

It  requires  five  men  to  successfully  operate  the  apparatus; 
one  hoist  man,  one  man  on  the  entry  to  stop  and  trim  the  cars, 
two  at  the  face  and  one  timberman.  The  foremost  saving  is 
in  the  loading  of  cars. 

A  full  crew  for  the  Myers- "Whaley  machine  consists  of  one 
runner  and  one  car  coupler,  who  load  as  much  as  15  to  20 
hand  shovelers.  With  these  machines  the  same  working  places 
can  be  loaded  out  twice  a  day,  instead  of  once  in  two  days, 
as  is  commonly  the  case.  Hence  a  machine-equipped  mine  will 
produce  a  given  tonnage  with  one  fourth  the  development 
required  where  the  loading  is  by  hand.  This  concentration, 
with  the  consequent  reduction  of  trackage,  ventilation  and 
mine  car  equipment,  effects  important  economies. 

The  machine  is  operated  by  one  man,  will  work  on  elec- 
tricity or  compressed  air;  and  is  self  propelling  forward  or 
backward.  The  machine  runs  on  either. 

The  machines  operate  at  the  rate  of  13  to  18  strokes  per 
minute,  with  a  power  consumption  of  .22  kw.-hr.  per  ton  loaded. 
Three  sizes  are  built  for  underground  mining  as  follows : 


Height 

Length 

Width 

Net  Weight 

Mine  Height 
Required 

No.  2.  46  in. 
No.  3.  47  in. 
No.  4.  54  in. 

19  to  21  ft. 
20  to  22  ft. 
22  to  26  ft. 

4  ft.    9  in. 
4  ft.  11  in. 
5ft.    4Jin. 

9,000 
11,000 
18,000 

4|  ft. 
5ft. 
6ft. 

MINING  COSTS 


121 


The  machines  are  all  capable  of  loading  at  over  1  ton  per 
minute.  The  smallest  handle  30  to  45  tons  per  hour,  and  the 
larger  sizes  from  50  to  60  tons  per  hour,  actual  shoveling  time. 

A  test  of  one  of  the  earlier  makes  of  this  machine  was  made 
at  the  mines  of  the  United  States  Coal  and  Oil  Co.  at  Holden, 
W.  Va.,  about  1910,  which  is  of  interest.  The  coal  at  Holden 


LOADER'S  DAILY  REPOKT 


Sixth  day  of  official  test  run. 


Holden,  W.  Va.,  Aug.  31,  1910 


Time  Started  7:30  a.m. 
Time  Stopped  12:03  a.m. 
Total  Time  a.m.  4  hr.  33  min. 


Time  Started  12:33  p.m. 
Time  Stopped  5:32  p.m. 
Total  Time  p.m.  4  hr.  59  min. 


No.  of  Men  Working: 
a.m.  4 
p.m.  4 


Working  Place 

Time 
Loading 
and 
Shifting 
Cars 

Time 

Changing 
Machine 

Time 
Lost 

Total 
Time 
Consumed 

Cars 
Loaded 

No   6  room            

1  h 

3  m 

1  h      1m 

2  h      4m 

24  tons 

No   7  room                

2  h    29  m 

25  m 

23  m 

3  h    17  m 

54  tons 

No.  1  face  room  (neck)  . 

1  h.  45  m. 
1  h    25  m 

24  m. 
11  m 

26m. 
None 

2  h.  35  m. 
1  h    36  m 

33  tons 
39  tons 

Totals  

6  h.  39  m. 

1  h.     3m. 

1  h.  50  m. 

9  h.  32  m. 

150  tons 

REMARKS. — Time  lost  No.  6  room — loose  setscrew  on  conveyor  sprocket  tightened 
between  7:30  and  8:15  =  45  min.  Off  track  7  min.;  pulling  down  coal  9  min.  Total, 
1  hr.  1  min. 

Time  lost  No.  7  room — waiting  on  driver,  1  min.;  off  track,  5  min.;  pulling  down 
coal,  17  min.  Total,  23  min. 

Time  lost  No.  1  face  room  neck — off  track,  9  min.;  pulling  down  coal,  17  min.  Total, 
26  min. 

Time  lost  No.  1  room — none 

J.  B.  HAILE,  Machine  Runner. 
Report  by  W.  Whaley. 

is  a  peculiarly,  hard,  tough  bituminous  coal,  known  as  No.  2 
gaseous.  Its  coherent  qualities  make  it  somewhat  difficult  to 
shoot  down  and  nearly  as  hard  to  shovel  as  large  lumps  of 
limestone. 

During  six  days  of  its  operation  a  record  was  kept  of  all 
features  of  the  run,  cars  loaded,  time  required  to  load  and 
shift,  time  to  change  the  machine  and  time  lost  for  any  cause. 
During  the  six  days  the  machine  loaded  768  tons  of  coal,  the 
time  of  loading  and  shifting  cars  being  36  hr.,  6  min. ;  time 
changing  the  machine  4  hr.  21  min. ;  time  lost,  14  hr.  33  min. ; 
total  time,  55  hr.  The  machine  loaded  out  four  rooms  per  day 


122  COAL  MINING  COSTS 

and  part  of  the  lost  time  was  due  to  delay  in  shot  firing,  in 
waiting  on  smoke  to  clear  out  of  the  room  and  other  items  not 
chargeable  to  the  machine.  A  copy  of  daily  report  of  the  last 
day  of  this  test  run  is  shown  herewith. 

In  spite  of  the  delays  and  lost  time  the  machine  averaged 
128  tons  per  day.  The  lowest  day's  work  (due  to  lack  of  coal) 
was  90  tons,  and  the  highest  150  tons.  The  average  time  of 
loading  and  shifting  a  car  was  8.4  min.,  or  21.3  tons  per  hour. 

The  territory  in  which  the  machine  worked  consisted  of 
seven  rooms  recently  turned  off  the  airway  in  No.  5  mine. 
Much  of  the  work  was  done  on  curves.  The  rooms  ranged  in 
width  from  21  ft.  to  27  ft.  and  all  but  two  had  two  tracks. 
The  rooms  were  on  the  butt  of  the  coal  which  greatly  increased 
the  difficulty  of  shooting. 

The  company  had  no  difficulty  whatever  in  keeping  the 
machine  supplied  with  cars,  which  were  hauled  to  and  from  the 
side  track  on  the  entry  by  a  mule. 

The  crew  of  the  machine  consisted  of  four  men,  as  follows : 
One  machine  runner,  one  man  in  front,  and  two  men  to  handle 
cars  and  pick  slate. 

Mining  and  loading  machines. — The  Ingersoll-Rand  cutter 
and  loader  consists  of  a  powerful  air-driven  puncher,  mounted 
on  a  carriage  over  a  conveyor.  A  lever  is  used  to  elevate  or 
depress  the  puncher  pick,  a  second  lever  moves  it  from  rib  to 
rib  through  the  agency  of  a  small  air  engine,  while  a  third 
lever  moves  the  whole  puncher  together  with  its  truck  bodily 
toward  or  away  from  the  face,  the  conveyor  remaining  station- 
ary within  certain  limits,  or  accompanying  the  forward  or 
backward  motion  of  the  puncher  as  desired. 

The  puncher  itself  makes  about  160  strokes  per  minute, 
and  is  controlled  in  the  same  manner  as  the  ordinary  hand- 
operated  machine,  going  from  rib  to  rib  and  making  a  cut 
extending  the  entire  width  of  the  entry.  The  conveyor  is 
driven  by  a  separate  engine  suitably  controlled  by  a  stop  valve. 
The  mine  car  is  filled  evenly  by  moving  it  about  three  times 
during  the  loading  process. 

Making  the  undercut  is  the  longest  part  of  the  operation, 
requiring  from  25  to  40  min.,  according  to  the  hardness  of  the 
coal.  The  knocking-down  process  ordinarily  takes  place  faster 
than  cars  can  be  supplied.  When  it  is  completed  the  conveyor 


MINING  COSTS  123 

is  pulled  back  about  7  ft.,  a  section  of  track  is  laid  and  the 
conveyor  again  moved  forward  into  position. 

Two  set-ups  during  the  shift  constitute  a  10-ft.  advance  of 
the  entry.  It  is  estimated  that  under  suitable  operating  con- 
ditions, such  as  high  coal  of  the  quality  found  in  the  Pittsburgh 
seam,  that  the  machine  should  average  250  ft.  of  advance  per 
month  of  25  days,  working  day  shift  only,  the  night  shift 
being  reserved  for  the  laying  of  tracks,  piping,  etc.  Where 
conditions  have  been  suitable  and  the  machine  has  had  an 
adequate  supply  of  cars,  as  much  as  20  ft.  of  advance  has  been 
made  in  a  single  shift.  The  average  advance  in  the  entries 
of  the  Annabelle  mine  in  West  Virginia  where  several  of  these 
machines  have  been  in  operation  since  1913,  is  somewhat  over 
10  ft.  per  shift.  The  cuts  at  this  mine  average  10y2  to  11  ft. 
wide  and  6y2  to  7  ft.  high. 

The  Jeffrey  cutting  and  loading  machine  undercuts,  shears, 
breaks  down,  and  loads  the  coal  into  the  mine  cars.  No 
explosives  are  used. 

The  machine  stays  in  the  entry  until  driven  as  far  as  desired. 
It  is  fed  forward  in  a  pan  similar  to  a  breast  machine. 

The  feed  forward  is  7  ft.  Average  depth  of  cut  6  ft.  6  in., 
width  5  ft. 

The  time  required  to  make  the  cut  depends  upon  the  con- 
ditions and  nature  of  the  coal.  Ordinarily  the  sumping  cut 
takes  less  than  30  min. ;  the  open  cut  less  than  20  min.  Mov- 
ing sideways  to  next  cut,  3  to  4  min. 

The  Jeffrey  cutting  and  loading  machine  has  two  ver- 
tical shearing  chains  between  which  is  mounted  a  frame 
carrying  a  number  of  heavy  punching  picks.  This  frame  is 
readily  raised  and  lowered  by  the  operator,  who  can  cause  the 
picks  to  strike  at  any  height  he  desires.  The  coal  falls  onto  a 
conveyor  which  is  made  thin  enough  to  go  into  the  kerf  cut 
by  the  undercutting  chain.  This  conveyor  carries  the  coal  to 
the  rear  end  of  the  machine  and  dumps  it  into  a  second  con- 
veyor, which  is  mounted  so  that  it  can  swing  at  any  desired 
angle  to  the  machine. 

After  a  cut  is  made,  the  machine  is  moved  sideways  by 
means  of  a  rope  hitched  to  a  jack  at  the  opposite  rib  and  then 
another  cut  is  made.  When  slate  is  to  be  piled  up  at  the  side 
of  the  room,  the  rear  conveyor  is  swung  sideways,  the  slate 


124  COAL  MINING  COSTS 

rolled  onto  the  front  of  the  machine  and  gobbed  by  the  two 
conveyors. 

In  a  test  run  the  machine  required  an  average  of  13^2 
min.  to  make  a  cut  and  about  3  min.  to  move  the  machine  to 
the  next  cut,  or  about  16  or  17  min.  time  for  each  cut.  The 
coal  loaded  averaged  21  tons  per  hour,  although  where  roof 
conditions  were  good  the  machine  had  loaded  30  tons  an  hour 
in  the  same  time.  The  height  of  the  coal  is  5  ft.  8  in.  In 
another  district  an  average  of  60  tons  per  hour  was  made  for 
the  three  months  that  the  machine  was  in  operation.  In  the 
driving  of  an  11-ft.  entry  it  has  averaged  20  ft.  advance  per 
shift  of  8  hr.  under  rather  unfavorable  circumstances. 

The  crew  for  each  machine  consists  of  a  machine  runner, 
a  helper  and  a  driver.  If  the  slate  is  heavy  it  may  require  an 
extra  man  for  this  handling. 

At  the  Valier  mine  in  Illinois  these  machines  have  been 
introduced  to  accomplish  rapid  development  work,  promote 
safety  through  elimination  of  explosives  and  prevent  the  shat- 
tering action  of  explosives  on  the  ribs  and  roof  of  the  entry. 
The  cutting  in  this  mine  is  unusually  hard,  but  under  ordinary 
conditions  an  advance  of  as  much  as  150  ft.  per  week  was 
made  with  each  machine,  by  working  three  shifts.  On  single 
shifts  the  machines  make  a  general  average  of  300  ft.  per  month. 
With  regard  to  the  economy  of  the  use  of  these  machines, 
it  is  the  opinion  of  Mr.  Carl  Scholz,  general  manager  of  the 
company,  that  the  cost  of  installation  and  operation  will  be 
repaid  many  times  by  the  saving  in  timbering  because  the 
coal  is  not  affected  by  the  use  of  explosives  when  these  machines 
are  used.  The  roof  will  stand  much  better  than  it  does  when 
so  shattered  and  timbering  will  be  unnecessary  in  most  parts 
of  the  entries.  These  machines  are  being  used  on  the  main 
west  entries  where  a  possible  distance  of  3^  mi-  can  he  driven. 
After  an  experience  of  about  1  yr.  in  entry  driving  in  this 
mine,  a  difference  between  the  standing  qualities  of  these 
entries  and  those  driven  with  explosives  can  be  observed. 

The  O'Toole  machine  undercuts,  breaks  down  the  coal  and 
loads  it.  It  breaks  up  the  coal  completely  and  is  thus  designed 
specially  for  use  in  mines  where  the  production  of  lump  coal 
is  of  no  advantage  as  for  instance  where  the  mine  is  producing 
solely  for  coke  making  purposes.  The  machine  consists  of  a 


MINING  COSTS 


125 


ffi 


TjH 


O5 


O    (N    CO   CO    to 


2  2  3  3  8  g  g  S3  g  3 

rH  <N    r-l 


i 


O  »O 

,— I 

§  g 


00   (N    00   O   O 
CQ    CO   Tt*    iQ   »-H 


CO 


111 


•i-lCOi-lCOi— iCOrHCOi-l 

:  -i  i  4<  i  J,  A  4,  i  J, 


•9    5 


126  COAL  MINING  COSTS 

motor-propelled  truck,  an  oscillating  chain-driven  revolving 
head  carrying  the  cutter  bits  and  a  scraper  conveyor  for  load- 
ing the  coal  into  mine  cars  at  the  rear  of  the  machine. 

A  daily  record  of  the  machine,  being  the  average  of  several 
days  is  as  follows :  Distance  advanced  38  ft. ;  power  con- 
sumption, 164  kw. ;  tons  mined,  89 ;  average  time  of  operation, 
6  hr.  40  min.  These  tests  were  conducted  simultaneously  with 
experiments  on  exhausting  coal  from  the  mine  by  means  of  an 
air  blast,  for  which  purpose  a  Root  blower  was  connected  to 
a  large  spiral-riveted  pipe  and  operated  exhausting.  The  pipe 
extended  for  several  hundred  feet  into  the  mine  and  was  con- 
nected with  the  mining  machine.  An  attempt  was  thus  made 
to  draw  the  coal  out  of  the  mine  and  into  a  tank  over  the 
railroad  track.  This  attempt  was  highly  successful  in  so  far 
as  coal  removal  was  concerned,  but  in  its  passage  through  the 
pipe  the  material  was  reduced  to  dust.  Although  this  con- 
dition was  not  disadvantageous,  as  far  as  immediate  coking 
was  concerned,  the  coal  obtained  could  not  be  shipped  long 
distances  in  open-top  cars. 

Because  readings  were  being  taken  at  the  time  that  the 
experiment  was  being  conducted,  the  results  obtained  are  not 
a  true  indication  of  what  the  machine  can  do.  One  instance 
of  this  kind  is  the  record  for  one  day  in  which  154  tons  were 
mined  in  9  hr.  and  47  min.,  and  the  advance  was  66  ft.,  the 
average  consumption  being  120  kw.  The  passage  driven  was 
in  all  cases  approximately  10  ft.  wide  and  7  ft.  high. 

A  table  showing  some  of  the  work  performed  by  this 
machine,  the  time  consumed  and  the  cost  is  presented  here- 
with. These  figures  are  as  of  1915  and  would  have  to  be  recal- 
culated, using  a  new  basic  rate,  in  order  to  bring  them  up  to 
date  and  render  them  comparable  with  present  conditions  and 
prices. 

Blasting.— It  has  been  found  by  experiment  that  when  a 
hole  is  drilled  in  rock,  loaded  with  powder,  tamped,  and  dis- 
charged, the  space  in  the  rock  after  explosion  has  the  shape  of 
a  cone.  If,  therefore,  a  hole  is  put  in  the  face  of  an  excavation 
and  given  no  inclination  the  explosive  will  blow  out  the  tamp- 
ing or  at  best  will  break  out  a  small  cone  near  the  mouth  of  the 
hole.  The  reason  for  this  is  that  the  gases  in  trying  to  escape 
follow  the  line  of  least  resistance,  which,  in  this  case,  is  the  drill 


MINING  COSTS 


127 


hole,  and  the  result  is  a  blown-out  shot.  If  a  hole  a  &  is  put  in 
at  an  angle  of  60  deg.  to  the  face  x  y  in  the  plan,  Fig.  50,  the 
cone  a  b  c  would  be  broken  out  were  there  no  reaction.  The 
line  of  least  resistance  in  this  case  is  b  d.  Gas  expands  equally 
in  all  directions,  and  since  it  cannot  escape  except  toward  the 
free  face,  two  forces  come  into  action,  one  direct  along  the  line 
b  d  and  the  reactive  force  along  the  line  b  c.  The  resultant  of 
these  two  forces  is  represented  by  the  lines  b  e  and  b  /,  conse- 
quently the  cone  is  narrowed  to  a  base  e  f  instead  of  a  c.  The 


FIG.  50. — Sketch  demonstrating  theory  of  blasting. 

greater  the  angle  the  narrower  will  be  the  cone  base,  and  the 
more  chances  there  will  be  for  a  blown-out  shot. 

If  the  hole  is  6  ft.  deep  the  length  of  the  line  of  least  resistance 
will  be  sin  60°  X  6,  or  0.86  X  6  =  5.16  ft.  This  is  6  —  5.16 
=  0.84  less  in  depth  than  when  the  holes  is  put  in  at  right  angles 
to  the  face  and  will  in  all  probability  be  a  blown-out  shot  where 
a  safe  charge  of  powder  is  used. 

If  the  hole  a  g  is  6  ft.  long  and  put  in  at  an  inclination  of 
45  deg.  to  the  face,  the  cone  would  have  the  dimensions  a  g  li 
were  there  no  resistance.  In  this  case  there  is  not  so  much 
resistance  to  the  breaking  of  rock  owing  to  the  line  of  least 
resistance  g  i  being  less,  which  allows  the  reactive  forces  greater 


128  COAL  MINING  COSTS 

play.  The  resultant  of  the  force  to  the  left  of  the  line  g  Us  g  k, 
and  as  this  comes  well  inside  the  rib  q  x  and  the  length  of  the 
line  g  i  is  sin  45°  X  6  or  0.7  X  6  =  4.2  ft.,  the  chances  are  that 
unless  the  explosive  is  properly  proportioned  and  the  charge  is 
carefully  tamped  there  will  be  a  blown-out  shot. 

If  a  hole  a  I  6  ft.  long  is  drilled  at  an  angle  of  30  deg.  to 
the  face,  the  cone  in  plan  would  have  the  shape  aim.  In  this 
case  the  line  I  m  extends  into  solid  rock  beyond  the  free  face 
x  y,  and  therefore  no  rock  would  be  broken  past  p.  In  this 
case  the  resultant  I  o  comes  to  the  corner  of  the  rib  at  p  and 
the  rock  broken  would  have  the  shape  a  I  p.  The  length  of  the 
line  of  least  resistance  I  n  is  sin  30°  X  6  or  0.5  X  6  =  3  ft., 
or  just  half  the  length  of  the  hole.  Ordinarily  with  care  a  com- 
promise can  be  effected  between  30  deg.  and  45  deg.  where  more 
work  can  be  accomplished  with  the  same  quantity  of  powder, 
and  this  may  be  at  35  deg. 

The  best  method  of  breaking  down  coal  at  any  mine  by 
the  use  of  explosives  is  determined  by  experiments,  which  are 
to  be  carried  on  with  intelligent  observations. 

The  observations  will  include  the  thickness  of  the  coal  bed ; 
the  area  of  the  face  to  be  excavated;  whether  it  is  advisable 
to  remove  the  whole  bed  or  let  some  of  it  remain  in  place; 
whether  advantage  shall  be  taken  of  partings  if  any  exist ;  the 
direction  of  the  excavation  with  reference  to  the  cleatage,  and 
the  tightness  of  the  coal.  When  these  matters  have  been 
definitely  determined,  then  a  few  experimental  shots  will  show 
the  best  positions  for  pointing  the  holes,  their  probable  depth 
and  the  quantity  of  explosive  needed  for  economically  breaking 
down  the  coal. 

The  system  of  mining  followed,  whether  it  is  shooting  off 
the  solid,  shearing  a  loose  end,  or  in  the  middle,  or  undercut- 
ting, will  require  the  same  careful  observations,  although  the 
pointing  of  the  drill  holes  will  vary  somewhat  with  each  system. 

Dynamite. — The  meaning  of  the  grade  distinctions  or  "per 
cent  strength"  mark  on  dynamite  is  somewhat  of  a  puzzle  to 
many  consumers,  and  often  a  source  of  misunderstanding 
between  manufacturer  and  customer. 

Originally  a  40  per  cent  dynamite  meant  that  the  dynamite 
contained  40  per  cent  of  actual  nitroglycerin  by  weight,  bat 
as  modern  dynamites  do  not  always  contain  this  proportion 


MINING  COSTS  129 

as  marked,  a  short  description  of  the  modern  practice  in  grad- 
ing is  necessary.  A  slight  knowledge  of  the  history  of  the 
manufacture  of  high  explosives  may  also  help  to  explain  the 
situation. 

The  dynamites  in  use  at  present  were  originally  known  as 
"active-base"  dynamites  in  contra-distinction  to  the  kiesel- 
guhr  dynamites  which  had  an  inert  base.  Low  grade  dynamites 
were  made  a  good  many  years  ago,  and  are  still  a  standard 
product,  the  composition  of  which  is  from  5  to  20  per  cent  of 
nitroglycerin  absorbed  in  a  combination  of  sodium  nitrate, 
sulphur  and  coal  dust,  but  as  these  are  not  a  good  absorbent, 
a  mixture  of  wood  meal  and  nitrate  of  soda  is  used.  With 
these  two  ingredients,  dynamites  can  be  made  with  different 
proportions  of  absorbent  to  nitroglycerin,  so  that  explosives  con- 
taining as  much  as  75  per  cent  or  as  little  as  15  per  cent  of 
nitroglycerin  could  be  made,  worked,  packed  and  exploded. 

It  was  also  found  that  with  an  active  base  like  wood  meal 
and  nitrate  of  soda,  a  dynamite  having  only  40  per  cent  nitro- 
glycerin would  develop  as  much  power  or  more  than  a  75  per 
cent  kieselguhr  dynamite.  A  reasonably  definite  proportion 
of  wood  meal  to  nitrate  of  soda  existed  at  which  an  explosive 
was  not  so  wet  that  it  would  leak  nor  yet  so  dry  that  it  could 
not  be  " punched"  into  the  paper  shells. 

The  proportions  of  wood  meal  and  nitrate  were  changed 
to  accord  with  any  change  in  percentage  of  nitroglycerin. 
More  wood  meal  and  less  nitrate  were  used  when  the  absorbent 
was  to  retain  a  large  percentage  of  nitroglycerin.  More  nitrate 
and  less  wood  meal  or  wood  meal  of  less  capacity  for  absorption, 
like  fine-grained  sawdust,  were  used  when  making  a  dynamite 
with  a  lower  percentage  of  nitroglycerin. 

Using  these  three  ingredients  with  minute  proportions  of 
other  nonexplosive  substances  required  to  stabilize  the 
dynamite,  a  type  of  high  explosive  known  as  "straight  dyna- 
mite" is  made  which  when  thoroughly  incorporated  out  of 
well  dried  and  pulverized  ingredients,  constitutes  the  standard 
of  strength  against  which  all  other  dynamites  are  graded. 

When  other  explosive  substances  are  incorporated  into 
dynamites  they  increase  the  power  over  the  straight  dynamite 
and  it  is  then  necessary  to  reduce  the  amount  of  nitroglycerin 
and  otherwise  modify  the  formula  so  that  the  new  compound 


130 


COAL  MINING  COSTS 


will  develop  the  same  power  in  actual  work  as  the  standard 
dynamite. 

For  instance,  when  guncotton  is  dissolved  in  nitroglycerin 
it  makes  a  sticky  jelly-like  substance  which  when  added  to  the 
wood  meal  and  nitrate  of  soda  makes  an  explosive,  the  cartridges 
of  which  are  much  more  powerful  than  those  of  the  same  size  in 
which  nitroglycerin  alone  is  used. 

If  such  an  explosive  were  graded  according  to  its  actual 
content  of  nitroglycerin,  the  cartridges  would  be  so  much  more 
powerful  than  those  of  the  standard  grade  of  dynamite  that  it 
might  not  be  safe  to  use  in  work  where  the  blasters  were  accus- 
tomed to  using  that  standard  grade,  as  it  would  break  the  ma- 
terial too  fine  and  throw  it  too  far  and  perhaps  do  much  damage. 

When  other  active  ingredients  in  the  absorbent  were 
employed,  it  was  found  necessary  to  reduce  the  amount  of 
nitroglycerin  until  the  mixture  developed  the  same  strength 
as  the  straight  dynamite  nitroglycerin  by  which  it  was  graded. 

There  are  now  many  explosives  in  the  market  which  con- 
tain no  nitroglycerin  at  all,  some  of  them  being  equal  to  a 
40  per  cent  straight  nitroglycerin  dynamite,  and  these  are 
graded  against  the  straight  nitroglycerin  dynamite. 

The  following  table  shows  the  total  production  of  explosives 
in  pounds  in  the  United  States,  according  to  the  United  States 
Bureau  of  Mines,  and  the  amount  of  the  various  kinds  of 
explosives  used  in  Pennsylvania  to  produce  91,626,964  tons  of 
anthracite  coal  and  172,965,652  tons  of  bituminous  coal  in  1913 : 


Black 
Powder 

High 
Explosives 

Per- 

missibles 

All  mines  in  the  United  States  
Pennsylvania  anthracite  region  
Pennsylvania  bituminous  region  .... 

194,146,747 
44,001,660 
14,652,931 

241,682,364 
16,093,035 
696,162 

27,685,770 
3,323,645 
6,715,028 

Thus  it  can  be  seen  that  the  anthracite  region  used  almost 
as  much  "permissible"  explosive  in  proportion  to  coal  pro- 
duced as  the  bituminous  region.  The  presence  of  gas  in  suffi- 
cient quantity  to  be  detected  on  an  ordinary  safety  lamp  makes 
the  use  of  permissible  explosives  advisable,  if  not  absolutely 
•necessary,  to  prevent  explosions  being  caused  by  the  flame  of 


MINING  COSTS  131 

the  black  powder  so  commonly  used  for  blasting  coal.  The 
mining  of  anthracite  consumed  over  three  times  as  much  black 
powder  and  23  times  as  much  "high"  explosive  as  the  mining 
of  bituminous  coal,  despite  the  fact  that  only  half  as  much  coal 
was  produced. 

The  figures  given  are  probably  reasonably  correct,  though 
the  operators  only  know  what  powder  they  sell  to  their 
employees  and  keep  no  track  of  what  is  purchased  from  other 
sources.  However,  the  amount  so  purchased  is  probably  not 
large.  Of  course  the  larger  consumption  of  powder  partly 
arises  from  the  fact  that  the  anthracite  miner  "lets  powder 
do  the  work."  If  the  bituminous  operator  invariably  shot  his 
coal  off  the  solid,  he  could  pay  his  pick  miners  a  lower  rate 
per  ton  and  would  nevertheless  pay  them  more  per  day  if  he 
could  only  sell  the  product.  The  greater  consumption  of  pow- 
der in  the  anthracite  region  is  therefore  not  without  its 
advantages  in  the  production  of  cheap,  though  less  marketable, 
coal. 

Hydraulic  cartridg'es. — A  hydraulic  cartridge  was  intro- 
duced about  1909  for  breaking  down  the  coal.  The  cartridge 
was  operated-  by  an  improved  form  of  pump.  Though  small 
it  is  extremely  powerful,  being  capable  of  exerting  a  pressure 
of  7  or  8  tons  to  the  square  inch.  It  is  of  special  design  and 
is  attached  directly  to  the  rigid  pipe  with  which  it  is  con- 
nected to  the  cartridge.  No  stand  is  required  so  that  the 
appliance  can  be  used  either  for  breaking  down  or  lifting 
up  the  coal  without  special  connections  and  can  be  operated 
in  holes  drilled  at  any  angle.  The  pump  is  fitted  with  a  water 
tank,  about  iy2  pints  being  required  for  the  operation,  but 
most  of  this  returns  to  the  tank  and  can  be  used  again.  A 
pressure  of  3  tons  per  square  inch  is  usually  required  in  seams 
up  to  4  ft.  thick,  and  this  gives  a  total  pressure  of  60,  90  and 
150  tons,  respectively,  on  the  three  sizes  of  cartridges  made. 

After  the  coal  has  been  undercut,  a  hole  S1^  in.  in  diameter 
is  drilled  into  the  coal,  slightly  less  deep  than  depth  of  holing. 
This  is  done  by  means  of  ordinary  machine  and  a  special  drill. 
The  hole  is  put  in  parallel  with  the  roof,  and  as  nearly  as 
possible  along  the  parting  to  which  the  coal  ordinarily  comes 
off,  or  just  below  it. 


132  COAL  MINING  COSTS. 

The  effect  of  the  use  of  this  machine  upon  the  working 
cost  is  slight,  while  its  general  advantageous  effect  upon  the 
selling  price  of  the  coal  is  quite  striking.  In  a  seam  using  five 
hydraulic  cartridges,  450  tons  of  coal  are  produced  per  day, 
of  which  75  per  cent  is  large  coal,  and  25  per  cent  small.  If 
in  the  same  seam  the  coal  is  brought  down  by  explosives,  the 
percentage  of  large  coal  decreases  to  about  65  per  cent,  and 
that  of  small  increases  to  about  35  per  cent.  The  profit  obtained 
by  use  of  cartridges  in  the  above  seam  on  450  tons  is  about 
$71  per  day. 

The  advantages  claimed  for  hydraulic  cartridges  are  greater 
percentage  of  lump  coal;  better  quality  of  coal  because  not 
shattered ;  better  quality  of  slack ;  no  waste  coal ;  no  dust ;  no 
flame ;  no  smoke ;  no  danger ;  the  work  is  done  in  the  daytime ; 
hence,  better  supervision  and  no  loss  of  work  in  waiting  for 
night ;  the  roof  is  not  broken  into  or  shaken ;  the  same  apparatus 
is  used  over  and  over  again. 

As  many  as  50,000  insertions  of  the  mining  cartridge  per 
annum  are  made  at  the  mines  of  the  Hulton  Colliery  Co.,  in 
England,  and  the  product  is  in  better  condition  as  it  leaves 
the  mine.  In  one  seam  alone,  the  Arley,  28,500  hydraulic 
"thrusts"  are  made  per  annum,  by  which  it  is  estimated  that 
92,600  tons  of  coal  are  produced. 

The  coal  face  is  well  supported  to  begin  with,  by  means  of 
sprags,  the  holing  being  made  to  a  depth  of  3  ft.  6  in.  or  4  ft., 
the  drill  holes  being  bored  a  few  inches  less  deep,  S1/^  in.  in 
diameter,  and  placed  at  intervals  of  6  ft.  along  the  entire  face. 
The  boring  of  each  of  these  holes  occupies  generally  about 
15  min.  After  boring  the  first  hole,  or  whenever  the  collier 
is  ready,  the  apparatus  is  inserted  into  the  hole  and  the  pump  is 
fixed  upon  a  movable  and  adjustable  support.  A  small  hand 
lever  is  first  actuated  until  pressure  is  reached  when  an  exten- 
sion handle  is  attached.  The  pressure  being  fully  on,  the 
enormous  power  of  the  apparatus  is  soon  apparent,  for  the 
coal  is  heard  to  be  rumbling  and  cracking.  This  is  allowed 
to  continue  until  the  back  portion  of  the  coal  is  broken  off, 
after  which  the  sprags  are  slightly  slackened.  By  a  continu- 
ance of  the  pumping,  the  pressure  is  brought  to  bear  at  the 
front  of  the  face  and  continues  to  spread  until  the  operation 


MINING  COSTS  133 

is  completed.  The  sprags  are  then  knocked  out,  whereupon 
the  whole  bank  of  coal  falls  down  in  large  pieces. 

Careful  tests  made,  both  with  explosives  and  hydraulic 
cartridges,  show  that  a  great  gain  in  lump  coal  is  obtained 
when  hydraulic  cartridges  are  used.  For  two  years  the  exact 
slack  and  round  percentages  were  taken  while  explosives  were 
used  and  the  result  showed: 

(a)  51  per  cent  large  lump;  17.3  per  cent  small  lump;  31.7 
per  cent  slack. 

Hydraulic  cartridges  were  then  introduced,  operating  over 
an  area  covering  one-half  of  the  same  mine,  explosives  being 
left  in  the  other  half.  For  the  first  12  months  after  using  the 
cartridges  the  percentages  were: 

(&)  55  per  cent  lump  coal;  18.5  per  cent  small  lump;  26.5 
per  cent  slack. 

Slightly  over  26  per  cent  slack  was  produced  therefore,  for 
the  whole  of  the  mine. 

Three  separate  tests  were  then  taken  of  cartridge  coal,  only, 
giving  an  average  of: 

(c)  64.37  per  cent  lump  coal;  13.87  per  cent  small  lump; 
21.76  per  cent  slack. 

The  commercial  value  of  the  hydraulic  cartridge  as  against 
explosives  on  the  basis  of  the  above  percentages  (a)  and  (c), 
and  taking  into  account  the  cost  of  using  the  apparatus  in 
England,  1907,  was  as  follows: 


HYDRAULIC  CARTRIDGE 

Working  Cost:  £       s.  d. 

Operator,  6  days  at  6s.  per  day 1      16  0 

Allow  for  charge  on  first  cost  (£30)  and  depreciation  of 

cartridges,  say 6  0 


Value  of  coal  produced  in  a  seam  using  5  cartridges: 
30  "thrusts"  each  per  day  produce  450  tons  of  coal. 
450  tons  at  78.24  per  cent  lump  and  small  lumps =352  tons 

at  10s 176        0      0 

450  tons  at  21.76  per  cent  slack =97.92  tons  at  5s 24        9      6 


200        9      6 


134  COAL  MINING  COSTS 

EXPLOSIVES 
Working  cost: 

On  the  same  face  a  shot  lighter  would  be  required,  but  as 
he  could  probably  fire  the  shots  in  less  time  this  amount 
is  deducted. 

Operator,  3  days  at  6s.  per  day 18      0 

Cost  of  explosives,  150  shots  at  4d 2       10      0 


380 
Value  of  coal  produced: 

450  tons  at  68.3  per  cent  lump  and  small  lump  =  307.35  tons 

at  10s 163       13      6 

450  tons  at  31.7  per  cent  slack  =  142.65  tons  at  5s 35       13      3 


Hydraulic  cartridge  gain  per  week :  199        6      9 

Less  cost  of  working;   £3  8s.  Od.  less  £2  2s.  Od 1        6      0 

Increased  value  of  coal  £200  9s.  6d. 
1896    9 


£  11  2s.  9d.  per  day 
Per  week  of  5  days.  .  . 55       13      9 


Extra  value  of  cartridges  per  week 56       19      9 

Extra  value  of  cartridges  per  year 2963        7      0 

At  this  colliery  15  cartridges  are  in  daily  use  with  equal 
success,  thus  trebling  the  advantage. 

The  price  of  complete  cartridge  was  £30  (f.o.b.  Liverpool). 

A  new  machine  of  this  kind  was  introduced  about  1920. 
Briefly  stated,  the  principle  employed  in  this  machine  is  a 
duplication  of  a  natural  process  found  in  every  mine,  in  that 
the  coal  is  broken  down  by  developing  and  applying  what 
might  be  termed  an  "artificial  squeeze. " 

Eectangular  incisions  are  cut  in  the  body  of  the  coal,  one 
near  each  rib,  and  sometimes  one  in  the  center  of  the  room. 
These  are  made  parallel  with  and  as  near  the  roof  as  possible, 
those  near  the  rib  being  cut  parallel  to  it  and  close  to  the 
corner  of  the  room.  Where  the  coal  has  the  usual  cleavage 
planes  and  slips,  the  center  incision  is  not  necessary,  as  the 
coal  in  that  case  will  break  down  readily  from  rib  to  rib.  The 
machine  used  for  cutting  these  incisions,  or  slots,  is  self-con- 
tained and  self-propelled. 

Sumping  is  the  only  work  required  of  this  bar — that  is,  it 
is  simply  inserted  and  withdrawn,  cutting  the  incision  in  less 


MINING  COSTS  135 

than  3  min.  after  the  machine  is  locked  in  position.  The  slots 
are  cut  to  approximately  the  same  depth  as  the  undercut.  The 
standard  height  of  kerf  is  4%  in.,  and  by  simply  changing  the 
width  of  the  cutter  bar  its  width  is  varied  to  suit  conditions. 
On  the  standard  machine,  cutter  bars  18,  24  or  32  in.  wide  can 
be  used. 

This  slotting  machine  is  simply  the  combination  of  two  old 
and  highly  perfected  devices — the  chain-track  type  of  tractor, 
and  the  undercutting  machine,  simplified  and  modified,  for 
cutting  an  incision  near  the  roof  instead  of  underneath  the 
coal.  Under  normal  conditions  of  operation  in  a  6-ft.  bed  three 
men  will  cut  in  eight  hours  on  the  average  the  required  num- 
ber of  slots  for  breaking  down  250  to  300  tons  of  coal. 

This  folding  steel  tubing  is  practically  indestructible  and 
solves  one  of  the  big  problems  in  this  process  of  mining,  as  it 
permits  of  water  being  conveyed  at  extreme  pressure  from 
the  pump  to  the  bar.  A  24-ft.  length  of  this  tubing  weighs 
approximately  42  Ib.  and  will  withstand  a  water  pressure  of 
30,000  Ib.  per  square  inch.  The  normal  pressure  employed  is 
10,000  Ib.  per  square  inch. 

The  entire  equipment,  however,  is  designed  for  using  water 
at  15,000  Ib.  per  square  inch,  if  such  a  pressure  is  required. 
All  parts  have  an  ample  factor  of  safety.  The  pump  is  of 
standard  design,  suitable  for  delivering  a  constant  volume  of 
water,  and  the  bars  are  of  proper  size  for  developing  the  neces- 
sary expansive  forces. 

The  expanding  bars  are  furnished  in  four  sizes.  The  smallest 
develops  a  total  expansive  force  of  1,000,000  Ib.  and  the  largest 
a  force  of  2,500,000.  The  pistons  have  large  flat  bearing  sur- 
face to  prevent  indentation.  The  thickness  of  bed,  character 
of  coal  and  other  local  conditions  determine  the  size  of  bar 
to  be  used.  Under  normal  conditions  the  bar  developing 
1,500,000  Ib.  of  expansive  force  unfailingly  will  break  down 
coal  9  to  10  ft.  in  thickness. 

The  advantages  resulting  from  the  elimination  of  the  use 
of  explosives  in  coal  mining  is  a  subject  on  which  volumes 
could  be  written.  It  effects  economies  in  production  and 
operation  that  are  not  anticipated. 

The  saving  in  life  and  property,  the  prevention  of  accidents 
in  general  and  other  humane  and  altruistic  features  are 


136  COAL  MINING  COSTS 

apparent.  The  roof  fall  is  the  greatest  danger  encountered  in 
coal  mining.  Eliminate  the  shattering  effect  of  explosives  upon 
the  roof  strata  and  the  accidents  will  be  greatly  reduced.  Also 
remove  the  powder  smoke  and  fumes  and  the  working  con- 
ditions become  more  healthful  and  pleasant. 

Economically,  the  use  of  powder  affects  every  item  of  pro- 
duction cost.  Less  timbering  is  required.  What  timber  is 
placed  is  never  blown  out.  Thus  losses  in  output  are  avoided 
and  the  cost  of  cleaning  up  the  resulting  roof  fall  is  eliminated. 
A  saving  in  costs  is  effected  through  the  elimination  of  shot- 
firers  and  other  highly  paid  labor.  No  expense  is  incurred 
for  installing  and  maintaining  shotfiring  apparatus  and  equip- 
ment. 

Shot  firing. — A  system  of  shot  firing  by  which  all  the  shots 
can  be  exploded  when  every  man  is  out  of  the  mine,  which  is 
not  expensive  in  installation  and  operation,  and  which  will 
not  heavily  restrict  the  output  of  the  mine  was  employed  by 
the  Utah  Fuel  Co.  at  its  mines  at  Sunnyside  and  Castle  Gate, 
Carbon  County,  Utah,  about  1908.  Here  the  coal  veins  vary 
from  5  ft.  to  10  ft.  in  thickness.  These  mines  are  operated  on 
room-and-pillar  system  and  have  a  dip  averaging  10  per  cent 
(or  5  deg.  and  45  min.).  At  Castle  Gate  the  main  haulage  is 
in  favor  of  the  load,  at  Sunnyside  No.  1,  against  the  load,  and 
at  Sunnyside  No.  2,  the  haulage  is  nearly  level  but  slightly  in 
favor  of  the  load.  All  these  mines  are  equipped  with  exhaust 
fans.  At  these  mines  all  shots  are  fired  from  the  surface  and 
not  then  until  it  is  known  positively  that  every  man  is  out  ol 
the  mine.  This  method  was  a  practical  working  success  and 
has  been  for  several  years  in  operation  in  mines  with  a  daily 
output  of  900  to  1400  tons  in  an  8-hr,  shift  and  can  just  as 
readily  be  adapted  to  a  mine  with  an  output  of  3000  tons  in  an 
8-hr,  shift. 

Two  rubber-covered,  or  weatherproof  (preferably  the 
former)  wires,  size  No.  6,  are  strung  from  the  power  plant  to 
the  mouth  of  the  mine  manway,  or  some  opening  through 
which  they  can  be  run  without  danger  of  disturbance  from 
wrecking  by  haulage  devices,  etc.  These  wires  act  as  feeders 
and  this  size  should  be  carried  into  the  mine  through  some 
opening  which  is  well  kept  up  and  at  the  same  time  approxi- 
mately divides  the  working  places  into  halves  in  order  to  reach 


MINING  COSTS  137 

all  these  working  places  with  the  minimum  expenditure  of  wire. 
From  this  feed  wire,  No.  12  rubber-covered  wire  is  run,  prefer- 
ably along  the  haulage  road  because  these  passageways  are 
kept  in  best  condition.  If  some  parallel  gallery,  however,  is 
kept  up  as  a  manway,  it  should  be  used  by  all  means.  But  if 
the  haulageway  is  used  the  wire  can  be  fastened  as  far  as  pos- 
sible from  actual  haulage  space.  From  this  No.  12  wire,  No. 
14  rubber-covered  wire  is  run  into  rooms,  pillars,  workings, 
etc.,  to  the  working  faces,  enough  wire  being  left  to  comfortably 
reach  every  point  in  which  a  shot  may  be  placed  without  being 
compelled  to  tighten  the  last  25  or  30  ft.  of  wire.  These  wires 
are  fastened  to  plugs  which  have  been  driven  into  holes  drilled 
into  roof  or  rib,  to  props,  to  cap  pieces,  to  legs  or  caps  of 
timber  sets,  etc.  They  are  attached  to  the  above  by  either 
two-wire  or  three-wire  porcelain  cleats  at  intervals  of  about 
25  ft.  or  sufficiently  often  to  insure  separation  of  the  wires  as 
the  insulating  covering  decays  and  falls  off  in  time. 

Giant  powder  is  used  to  bring  down  the  coal,  the  per  cent 
of  nitro glycerin  depending  on  the  hardness  and  tenacity  of 
the  coal  and  on  the  use  to  which  the  coal  is  to  be  put.  Sunny- 
side  coal  is  almost  wholly  used  for  coking,  and  crushing  being 
necessary,  the  more  fines  which  can  be  produced  in  the  mine 
the  better  after  eliminating  considerations  of  safety  to  roof, 
etc.  As  Castle  Gate  coal  is  sold  commercially,  only  such 
strength  of  powder  is  used  as  will  bring  down  the  coal  with 
a  minimum  of  fines.  "Reliable  exploders"  are  used  to  dis- 
charge the  shots,  the  exploder  being  merely  laid  upon  or  tied 
to  the  powder  with  a  half  hitch  of  the  two  wires,  about  6  ft. 
of  which  are  attached  to  each  exploder.  These  wires  are 
attached  to  the  No.  14  rubber-covered  wires,  previously  men- 
tioned, in  parallel  and  not  in  series.  If  attached  in  series,  and 
several  shots  in  the  series,  a  defective  exploder,  or  exploder 
wire,  may  cause  several  missed  shots,  but  when  placed  parallel, 
a  defective  exploder  will  affect  only  the  hole  in  which  it  is 
placed.  In  making  wire  connections  all  wires  should  be  bared 
and  twisting  should  be  resorted  to  in  order  to  insure  good 
contact. 

The  power  used  at  Sunnyside  is  derived  from  a  continuous- 
current  generator  giving  500  volts,  the  cost  of  power  for  shot 
firing  being  practically  nothing,  as  the  current  is  used  for  but 


138  COAL  MINING  COSTS 

a  few  minutes  daily,  while  for  the  remainder  of  the  time  it  is 
used  for  underground  hoists,  electric  locomotives,  and  light- 
ing. If  lack  of  electric  power  is  the  only  consideration  in  the 
way  of  installing  an  electric-shooting  system,  a  small  plant 
can  be  installed  and  operated  at  a  reasonable  cost  and  the 
surplus  power  can  be  most  profitably  devoted  to  electric  haul- 
age, or  to  underground  or  surface-lighting  systems. 

The  cost  of  installing  the  firing  system  with  electric  power 
already  at  hand  is  trivial.  At  Sunnyside  Mine  No.  1  the  com- 
plete cost  of  installation  from  power  house  to  the  working  faces, 
including  new  material  for  the  entire  line,  was  but  $1250,  of 
which  $850  was  for  material.  This  material  included  6000  ft. 
of  No.  6  rubber-covered  wire,  26,000  ft.  of  No.  12  rubber- 
covered  wire,  and  30,000  ft.  of  No.  14  rubber-covered  wire; 
also  10  switches,  about  twenty  16-ft.  poles  with  cross-arms, 
etc.,  for  the  surface  line,  insulators,  porcelain  cleats,  etc.  If  a 
mine  produces  1000  tons  of  coal  per  day  for  250  days  per  year 
or  a  total  of  250,000  tons  per  year,  the  total  cost  of  the  system 
if  entirely  paid  for  the  first  year  would  amount  to  but  one-half 
cent  per  ton.  In  operation,  a  checkman  at  $75  per  month, 
shot  firer,  or  wireman,  at  $3.25  per  day,  and  an  inspector  at 
$100  per  month  would  amount  to  $2912.50  per  year.  On  an 
output  of  250,000  tons  this  would  be  but  about  Ic.  per  ton, 
leaving  out  material  for  repairs  as  of  no  consequence. 

The  system  was  installed  at  Sunnyside  Mine  No.  2  which  is 
as  extensive  as  No.  1  mine,  for  $640,  the  difference  being  due 
to  the  fact  that  the  poles,  etc.,  were  utilized,  which  were  already 
installed  for  the  electric-haulage  system. 

Possible  decrease  of  output  due  to  missed  shots,  which 
prevent  miners  from  working,  may  be  advanced  as  a  reason  for 
condemning  the  system.  This  argument,  however,  is  disposed 
of  by  the  fact  that  out  of  an  average  of  300  shots  fired  daily, 
but  1  per  cent  missed  except  those  occasionally  due  to  the  fall- 
ing of  roof  on  feed-wires.  The  shots  of  a  whole  level  have 
been  known  to  miss,  but  this  occurrence  has  been  very  rare. 
At  any  rate  no  appreciable  diminution  of  output  was  noticed 
when  the  change  of  shot-firing  system  was  made  at  the  Sunny- 
side  mines. 

To  offset  the  cost  of  installation,  operation,  and  of  possible 
diminution  of  output,  may  be  advanced  the  fact  that  in  the 


MINING  COSTS  139 

5  yr.  the  system  has  been  in  operation  at  Sunnyside  where 
over  500  miners  have  been  continually  employed,  not  one 
accident,  however  trivial,  has  been  caused  by  shot  firing 
directly  or  indirectly.  Under  ordinary  methods,  it  would  be 
expected  that  several  accidents  due  to  shooting  would  have 
occurred  in  this  length  of  time,  any  one  of  them  resulting  in 
damage  suits  involving  $10,000  or  more.  The  loss  of  one  suit 
would  amount  to  practically  as  much  as  the  operating  cost  of 
the  system  for  4  yr. 

Daymen. — The  price  paid  the  miners  and  loaders  for  plac- 
ing the  coal  on  the  mine  car  at  the  working  face  is  the  largest 
single  item  in  the  costs  of  operating  bituminous  coal  mines. 
The  next  largest  is  the  day-labor  expense.  The  former  is  prac- 
tically fixed  for  each  region  and  does  not  vary.  Every  ton  of 
coal  produced  costs  just  so  much,  and  regardless  of  the  manage- 
ment it  remains  the  same. 

The  superintendent's  ability  to  handle  and  operate  a  mine 
successfully  hinges  principally  on  the  one  item — the  day-labor 
cost  of  operation.  This  he  can  reduce  or  raise  at  will  within 
certain  limits.  He  can  improve  the  underground  conditions 
while  better  market  prices  rule  or  reduce  the  amount  of  labor 
done  during  dull  times.  The  success  or  failure  of  the  mine 
depends  largely  on  this  labor  cost.  It  is  subject  to  more  varia- 
tion than  any  other  cost  item  on  account  of  its  being  dependent 
almost  entirely  upon  the  individual  ability  of  the  management, 
and  this  quality  has  as  many  variations  as  there  are  individuals. 

The  question  of  what  this  cost  should  be  is  much  disputed, 
and  many  estimates  are  made  based  on  any  one  mine  or  group 
of  mines.  Each  superintendent  knows  what  he  is  doing  him- 
self, but  there  is  more  or  less  reticence  in  exchanging  figures 
with  neighboring  mines. 

An  average  value  for  this  cost  of  labor  is  important,  and 
the  impracticability  of  obtaining  average  figures  over  large 
sections  for  periods  of  considerable  time  has  led  to  much  mis- 
understanding and  criticism  of  the  local  management.  The 
average  cost  is  not  what,  in  the  opinion  of  a  few  men,  the  work 
should  be  done  for,  nor  is  it  a  few  months'  test  run.  It  is  a 
figure  covering  a  long  period  of  time,  allowing  for  considerable 
mining  difficulties  and  embodying  some  allowance  for  the  labor 
necessary  to  the  general  maintenance  and  repairs  of  the  mine 


140 


COAL  MINING  COSTS 


and  plant.  It  should  include  a  wide  variety  of  conditions,  both 
favorable  and  adverse.  Granting  that  this  average  can  be 
established,  the  benefit  to  the  operators  is  evident. 

The  accompanying  diagram,  Fig.  51,  is  compiled  from  pub- 
lished figures  of  454  mines  in  Pennsylvania  and  West  Virginia. 
The  number  of  mines  entering  into  the  average  is  so  great 
that  any  inaccuracy  that  may  occur  in  the  report  from  any 
one  mine  would  be  inappreciable  in  the  result.  Any  temporary 
conditions  such  as  an  unusually  high  or  low  cost  simply  in- 
fluence the  result  in  making  the  figures  represent  average  con- 


Da'i\\(      Capacities 

400  to  600  to 

600  Tons          800  Tons 


800  to 
1000  Tons 


Over 
1000  Tons 


Small  circled  figures  — 
represent  trie  number  of— 
mines  entering  into  the— 
average 


FIG.  51. — Diagram  showing  labor  cost  at  454  Pennsylvania  and  West  Virginia 

coal  mines. 


ditions  through  long-time  operation.  Also  mines  were  selected 
which  work  over  230  days  in  the  year,  so  that  the  estimate 
may  be  taken  as  of  mines  producing  their  capacity  and  main- 
taining their  equipment  and  mine  conditions. 

The  foundation  of  the  diagram  is  the  mean  force  of  day 
employees  requisite  at  the  mine  for  a  period  of  one  year;  and 
the  average  daily  shipments  are  based  on  the  annual  produc- 
tion. Some  mines  included  in  the  average  may  be  going  through 
that  period  of  reconstruction  and  betterment  that  is  advisable 
periodically.  Others  may  be  using  every  means  of  economy. 
In  fact,  the  figures  may  be  taken  as  being  very  fair  operating 


MINING  COSTS  141 

expenses  covering  a  considerable  period  of  time  and  variable 
conditions.  They  do  not  represent  the  cheapest  mine  nor  the 
most  expensive.  Certain  mines  must  necessarily  call  for  higher 
labor  costs  than  others,  but  these  conditions  are  known,  and 
any  such  allowance  made  to  fit  the  individual  case.  Although 
the  figures  include  an  average  amount  of  rock  work  they  do 
not  represent  exceptions ;  and  they  represent  the  average  haul 
over  the  large  number  of  mines. 

The  basis  of  the  diagram  is  the  number  of  tons  output  per 
day-employee  or  day-laborer  per  working  day.  It  does  not 
take  into  consideration  any  extra-time  track  men,  drivers  and 
others  who  may  be  working  while  the  mine  is  idle.  It  is 
customary  for  some  day-men  to  work  in  idle  time  on  something 
that  may  be  then  more  easily  done.  Although  this  is  a  small 
percentage  of  the  regular  day-labor  expenses  it  is  something, 
and  it  would  accordingly  increase  the  cost  given  in  the  dia- 
gram. In  explanation  of  the  chart,  it  may  be  added  that  the 
force  of  miners  or  loaders  necessary  to  produce  the  tonnage  is 
not  figured  in  the  day-labor  cost,  the  coal  being  loaded  oh  the 
mine  car  at  the  working  face  by  the  requisite  number  of 
miners.  The  coal  is  hauled  by  drivers,  track  laid  and  repaired 
by  trackmen,  doors  opened  by  trappers ;  these,  with  the  motor- 
men,  brakemen,  dumpers,  trimmers,  car  shifters  and  all  other 
day-labor  necessary  to  move  this  tonnage  from  the  working 
face  to  the  loaded  railroad  car  are  represented  on  the  diagram 
in  the  " Production  in  tons  per  day-laborer  per  day*'  and 
11  Labor  cost  in  cents  per  ton,  estimating  the  average  day  wage 
at  $2.20  per  day." 

Firebosses  are  included  when  employed,  also  shopmen  and 
carpenters;  but  not  the  superintendent  or  mine  foreman  or 
the  mine-office  force.  It  is  simply  the  bare  day-labor  cost.  The 
cost  per  ton  is  derived  from  the  tons  per  day-man  per  day. 
The  greater  the  tonnage  the  less  the  cost.  The  averages  given 
in  the  lower  half  of  the  diagram  show  that  the  production  per 
day-man  runs  from  8.7  tons  for  the  low-capacity  machine  mines 
up  to  13.4  tons  for  the  low-capacity  pick  mines.  These  are 
the  extremes  in  the  averages. 

To  reduce  these  figures  to  cents  per  ton,  an  average  day 
wage  throughout  the  industry  must  be  assumed.  This  average 
varies  in  different  regions,  and  the  upper  half  of  the  diagram 


142  COAL  MINING  COSTS 

may  have  to  be  readjusted  to  a  small  extent  to  fit  the  prevail- 
ing wage  scale.  Curiously  the  average  production  per  day-man 
per  day  remains  practically  unchanged  throughout  many  thick- 
nesses and  characters  of  coal — for  the  central  Pennsylvania 
thin  steam  coals,  western  Pennsylvania  thick  gas  and  coking 
coals,  thick  steam  coals  of  southern  West  Virginia  and  the 
thick  and  thin  gas  coals  of  central  and  northern  West  Virginia. 

If  any  variation  from  this  is  to  be  noted  it  is  in  a  little 
greater  efficiency  in  the  regions  where  the  higher  wage  rates 
are  in  effect.  For  this  reason  a  moderate  day  rate  was  estab- 
lished; namely,  $2.20  per  day — little  higher  than  the  existing 
average  in  the  low-rate  regions  in  1915,  and  lower  than  those 
existing  in  western  Pennsylvania.  It  is  doubtful  if  the  costs 
per  ton  shown  on  the  diagram  can  be  changed  materially  for 
a  wide  range  of  mining  conditions  and  wage  rates  throughout 
the  bituminous  coal  regions. 

There  is  an  almost  inexplicable  variation  in  the  production 
per  day-laborer  per  day  when  individual  mines  are  considered, 
extreme  cases  running  from  20  tons  down  to  5  tons.  This 
indicates  some  confusion  in  this  mine-labor  item  and  the  lack 
of  any  well-founded  base  for  what  this  cost  really  should  be. 

It  will  be  noticed  that  51  pick  mines  with  a  capacity  of  less 
than  200  tons  daily  were  operated  with  a  labor  cost  of  16.6c. 
per  ton — the  cheapest  group  of  mines.  The  interpretation  is 
that  these  mines  are  so  small  and  simple  in  their  organization 
and  so  few  men  are  employed  that  it  is  almost  impossible  to 
blunder  in  their  operation.  The  only  explanation  advanced  as 
to  the  rise  in  costs  in  pick  mines  up  to  400  or  600  tons  capacity 
is  that  these  mines  outgrow  the  crude  management  of  the 
smaller  mines  up  to  that  point ;  and  from  there  up  to  the  1000- 
ton  mark  the  more  experienced  management  is  in  evidence. 

Small  capacity  machine  mines  are  a  failure,  as  the  cost 
of  the  power  house  and  appliances  are  prohibitive  for  such 
small  tonnages,  and  the  operating  costs  of  these  mines  decline 
generally  up  to  1000-ton  capacity.  There  is  little  variation  in 
this  decline,  as  their  equipment  calls  for  experienced  men  in 
the  start. 

The  reason  for  increase  in  cost  for  all  mines  over  1000 
tons  capacity  can  only  be  speculation,  but  the  evidence  points 


MINING  COSTS 


143 


to  their  being  too  large  for  almost  any  human  being  to  exercise 
the  proper  supervision.  Large  mines  are  almost  always 
equipped  with  the  most  modern  labor-saving  appliances,  and 
their  labor  costs  should  be  correspondingly  less. 

A  further  study  of  these  data  gives  the  following  results : 

1.  Averaged  42  tons  per  day  for  one  year  with  two  day  employees,  or 
21  tons  per  day-man  per  day. 

2.  Averaged  365  tons  per  day  for  one  year  with  16  day  employees,  or 
22.7  tons  per  day-man  per  day. 

3.  Averaged  429  tons  per  day  for  one  year  with  17  day  employees,  or 
25  tons  per  day-man  per  day. 

4.  Averaged  218  tons  per  day  for  one  year  with  eight  day  employees, 
or  27.2  tons  per  day-man  per  day. 

5.  Averaged  131  tons  per  day  for  one  year  with  five  day  employees,  or 
26.2  tons  per  day-man  per  day. 

The  average  of  these  is  24.4  tons  per  day-man  per  day,  and 
these  mines  were  selected  for  their  unusually  efficient  opera- 
tion. There  are  very  few  of  them  in  the  list  of  454  mines,  and 
they  are  given  to  show  that  this  grade  of  efficiency  has  been 
reached  by  the  willingness  of  the  employees.  The  aggregate 
production  and  the  number  of  day-men  employed  at  all  of  the 
454  mines  show  their  average  efficiency  to  be  11.4  tons  per  day- 
man per  day.  These  figures  and  the  following  tabulation  in- 
clude all  the  employees  inside  and  outside  the  mine: 


Tons  per  Day-Man 
per  Day 

Cost  per  Ton 
(Labor  Averaged 
at  $2.20  per  Day) 

Five  mines  

24.4 

$0  0902 

454  mines  

11  4 

0  1930 

In  the  454  mines  the  labor  work  cost  10.28c.  per  ton  more 
than  it  ought  to  have  cost  on  a  fair  valuation. 

The  saving  of  10.28c.  per  ton  is  a  minimum,  as  many  of  the 
454  mines  are  operated  with  a  very  efficient  force  of  men,  and 
they  include  the  five  mines. 


144 


COAL  MINING  COSTS 


Looking  at  the  question  from  another  point  of  view,  we 
have: 


Tons  per 
Day-Man 
per  Day 

Cost  per  Ton 
(Labor  Averaged 
at  $2.20  per  Day) 

Saving 

64  efficient  pick  mines  
100  inefficient  pick  mines 

16.76 
9  66 

$0.1313 
0  2277 

$0  0964 

139  efficient  machine  mines  .  .  . 
151  inefficient  machine  mines.. 

13.44 
8.29 

0.1637 
0.2654 

0.1017 

One  writer  has  suggested  a  " labor  factor,"  to  fix  the  ratio 
of  the  number  of  company  men  to  the  tonnage  produced.  This 
gives  certain  valuable  comparisons  for  the  immediate  vicinity 
where  the  labor  conditions  may  be  all  the  same,  but  is  subject 
to  fluctuations  in  comparisons  between  all  pick,  puncher,  short- 
wall  and  longwall  mining,  according  to  the  division  drawn 
between  miners  and  others. 

Efficiency  of  the  daymen  may  be  high,  but  the  miners  and 
loaders  may  be  waiting  for  cars  and  not  working  to  the  best 
advantage.  An  extra  employee  here  and  there  inside  may  not 
in  himself  be  working  to  the  best  of  his  time,  but  he  may  be 
helping  the  miners  to  double  their  output.  A  better  basis  of 
comparison  would  appear  to  be  one  involving  the  output  per 
man  employed  inside. 

The  accompanying  table  shows  a  comparison  between  the 
largest  producers  in  the  central  Pennsylvania  field  operating 
on  the  thinner  seams  of  coal,  the  Freeports  and  Kittannings, 
from  data  taken  from  the  state  mine  inspector's  report.  These 
figures  may  be  inaccurate  in  minor  details,  but  are  probably 
equally  correct  for  all.  The  collieries  are  arranged  according 
to  the  thickness  of  the  seam,  and  the  final  tonnage  is  divided 
by  the  thickness  of  the  seam  in  feet,  so  as  to  put  all  on  an 
even  basis. 

It  will  be  noticed  that  there  is  a  wide  variation  in  the  ton- 
nage handled  by  the  men  outside  the  mine.  This  is  to  bo 
expected,  as  conditions  vary  more  widely  outside  than  inside, 
the  matter  of  picking  probably  having  more  influence  than  any 
other.  Some  companies  purchase  power,  but  most  of  those 


MINING  COSTS 


145 


on  the  list  are  so  large  they  have  found  it  economical  to  pro- 
duce their  own  or  are  producing  it  because  their  plants  were 
installed  before  it  was  convenient  to  purchase  power.  All 
employees  engaged  in  the  manufacture  of  coke  have  been 
eliminated.  With  a  general  knowledge  of  the  conditions  exist- 
ing in  each  case,  a  fairly  reliable  comparison  should  be  reached. 

PRODUCTION  EFFICIENCY  OF  VARIOUS  COMPANIES  IN  CENTRAL  PENNSYLVANIA 
(Compiled  from  Mine  Inspector's  1912  Report) 


Thickness 
of  Coal, 
Inches 

Com- 
pany 

TONS  PER  MAN  PER  DAY 

Foot- 
Tons 

Standing 

Inside 

Outside 

Inside  and 
Outside 

38 
40 
42 

A 
B 
C 
D 
E 
F 
G 
H 
I 
J 
K 
L 
M 
N 
0 
P 

Q 

R 

2.27 
2.91 
3.60 
3.16 
2.50 
3.65 
3.40 
3.60 
3.39 
2.31 
3.95 
3.56 
3.70 
3.31 
4.68 
3.40 
4.12 
3.52 

0.72 
0.87 
1.28 
0.90 
0.71 
0.97 
0.89 
0.90 
0.85 
0.46 
0.84 
0.73 
0.74 
0.66 
0.85 
0.61 
0.68 
0.58 

12 
6 
1 
3 
13 
2 
5 
4 
7 
18 
9 
11 
10 
15 
8 
16 
14 
17 

3.45 
4.43 
3.56 
2.72 
3.96 
3.68 
4.06 
3.53 
2.92 
4.67 
4.73 
4.12 
4.15 
5.53 
3.90 
4.77 
4.27 

19.3 
21.3 
28.1 
30.4 
45.2 
43.3 
31.7 
87.0 
11.1 
25.8 
14.3 
36.5 
16.4 
30.3 
25.7 
34.4 
20.7 

45 
46 

48 

56 

58 
60 

66 

67 

72 

Average.  .  .  . 

30.1 

0.79 

The  United  States  Coal  &  Coke  Co.  in  1917  offered  prizes 
to  the  men  who  would  earn  the  most  money  in  the  loading  of 
coal  during  the  last  half  of  June,  the  results  of  which  are 
given  in  the  accompanying  table.  These  figures  are  of  interest 
chiefly  in  showing  the  maximum  labor  effort  that  can  be  real- 
ized, some  of  the  men  having  worked  day  and  night  and  laying 
off  for  a  full  week  after  the  test  period  was  over. 


146 


COAL  MINING  COSTS 


EARNINGS  OF  THE  THREE  BEST  COAL  LOADERS  AT  EACH  PLANT  OF  THE 

GARY  MINES 


Name 

No.  of 
Cars 

Rate 

Days 

Rate 

Slate 

Rate 

Total 
Earnings 

No.  2  Works: 
K  Kolony 

117 

$1  02 

3.0 

$2.55 

$126  99 

G  Belone 

103 

1  11 

13 

$0  84 

125  25 

L  Sabo   

95 

1.04 

2.0 

2.55 

10 

.84 

112.30 

No.  3  Works: 
J  Heedo 

114 

73 

2  4 

2  50 

89  33 

J  Seko 

96 

80 

18  0 

18 

55 

86  70 

S  Rinko  

/    47 

.85  \ 

8.0 

2.50 

3 

.55 

81.59 

No.  4  Works: 
P.  Clinchook... 
S  Miller 

\   27 

152 
150 

.73] 

1.11 
1  11 

1.0 

3.00 

3 
9 

.84 
84 

174.64 
174  06 

J  Kromba 

95 

1.11 

1.0 

2.55 

6 

.84 

113.04 

No.  5  Works: 
F  Lasos 

50 

1  37 

1.0 

2.55 

55 

.84 

117  25 

P  Blink 

64 

1  37 

1  0 

2  55 

29 

.84 

114  59 

S  Borsus  

51 

1.37 

1.0 

2.55 

50 

.84 

114.32 

No.  6  Works: 
T  Hajeck 

131 

1  11 

1  5 

2  55 

149  39 

J.  Sabinsky..  .  . 
A.  Ha  i  ser 

102 
83 

1.11 
1  11 

1.0 
0  6 

2.55 
2  55 



115.77 
93  86 

No.  7  Works: 
S  Kozlske 

96 

1  11 

0.3 

2.55 

107.41 

W.  Karbovich.  . 
F  Near 

87 
(    68 

1.04 
.97  \ 

3.3 

2.55 

8 

.36 

101.01 
100  28 

No.  8  Works: 
W.  Skarino.... 
B  Slovak 

\   33 

115 

62 

1.04J 

1.15 
1  05 

0.5 

2.55 

5 

28 

.84 
84 

137.87 
91  17 

Z  Rugan  

64 

1.11 

4.0 

2.55 

6 

.84 

86.28 

No.  9  Works: 
F.  Markangelo. 
D  Paron 

209 
194 

1.11 
1  11 

1.0 
1.6 

3.00 
2  55 

1 

.55 

235.54 
219.59 

S  Darboldi 

205 

1  11 

2.4 

2.55 

233.78 

No.  10  Works: 
N.  Carper  
A  Dodaney 

168 
96 

1.11 
1  11 

2.5 

8  0 

2.55 
2  55 

1 

.55 

193.55 
126  96 

M  Simms 

100 

85 

1  0 

2  55 

30 

.55 

104  05 

No.  11  Works: 
S.  Coimoin  .... 

P  Miller  

240 
247 

.95 
.95 

13.0 

/2.0 

2.50 
4.101 

1 
1 

.55 
.55 

261.05 
245.90 

P  Pope 

/    88 

.90\ 

\1.0 

2.50/ 

232  15 

No.  12  Works: 
J  Drake 

\161 
110 

.95/ 
1  11 

122  10 

J  Scolgal 

90 

1  11 

99  90 

E  Lester 

80 

1.11 

2.0 

2.50 

93.80 

MINING  COSTS 


147 


EARNINGS  OP  THE  THREE  BEST  COKE  PULLERS  AT  EACH  PLANT  OF  THE  GARY 

MINES 


Name 

Ovens 

Rate 

Ovens 

Rate 

Total 
Earnings 

No.  2  Works: 

41 

$1.55 

43 

$1  25 

$117  30 

D.  Hostin  

(   13 

1.55 

7 

1.251 

91  82 

J.  Hostin  

I  98 
(   12 

.1(4 

1.55 

17  Days  at 
10 

2.75J 
1.251 

83  98 

No.  3  Works: 
E  Young  

1293 
50 

.16* 
1.25 

1.8  Days  at 
17 

2.40/ 
1  55 

88  85 

W  Young         

38 

1  25 

15 

1  55 

70  75 

30 

1.25 

18 

1  55 

65  40 

No.  4  Works: 
P  Kellam  

20 

1  55 

56 

1  25 

101  00 

N.  Kellam  

16 

1.55 

36 

1  25 

69  80 

D  Kellam  

13 

1.55 

32 

1  25 

60  15 

No.  5  Works: 
S  Hodge  

19 

1  55 

49 

1  25 

90  70 

C.  Wade  

14 

1.55 

36 

1  25 

66  70 

L.  Nacne  

I13 

1.55 

25 

1.251 

54  60 

). 

1  3  Days  at 

2  40  / 

No.  6  Works: 
J.  Mills  

/     9 

1.55 

20 

1.251 

81  19 

T.  Mills  

1219 
f   13 

.16 
1.55 

3  Days  at 
37 

2.40J 
1.251 

70  93 

\ 

1  8  Days  at 

2  40  / 

J.  Moore  

I" 

1.55 

19 

1.251 

44  27 

\ 

1  4  Days  at 

2  40  / 

No.  7  Works: 
W.  Wilburn  

17 

1  55 

37 

1  25 

77  60 

J  Santos 

(  u 

1.55 

15 

1.251 

71  27 

E  B  Rose  

X... 
(  8 

1.55 

14.7  Days  at 
21 

2.40J 
1.251 

40  25 

I  ... 

0  6  Day  at 

2  40  / 

No.  8  Works: 
W  Kellam 

(     6 

1.00 

51 

1.251 

113  15 

G  Hall 

I... 
34 

1  25 

28 
12 

1.55J 
1  55 

61  10 

J  Martin 

25 

1  25 

14 

1  55 

53  60 

148 


COAL  MINING  COSTS 


The  outside  employees  used  in  the  anthracite  regions  would 
compare  better  with  those  required  in  the  bituminous  if  we 
were  to  eliminate  all  the  men  engaged  solely  in  the  cleaning 
and  sizing  of  the  coal.  Pennsylvania  anthracite  statistics  sup- 
ply information  in  which  " slate  pickers"  are  segregated  and 


TONNAGE  PER  EMPLOYEE  IN  PENNSYLVANIA,  WEST  VIRGINIA  AND  ILLINOIS 


Tons  per 

Tons  per 

Employee 

Employees 

Per  Cent 

Employee 

per 

per  Year 

Working 

Day 

Pennsylvania  anthracite,  1913*:. 

Miners  

78,319 

44.67 

1136f 

4.69 

Other  inside  employees  

50,348 

28.72 

1768f 

7.31 

Outside  employees  }  

46,643 

26.61 

1964§ 

8.12 

Total                   

175,310 

100.00 

523 

2.16 

*  All  mines  reporting. 

t  Based  on  all  tonnage  except  that  from  culm-banks.  The  whole  tonnage  was  91,626,964 
in  1913,  2,619,785  being  from  mines  and  strippings,  leaving  89,007,179  tons  obtained  from 
culm-banks. 

t  This  includes  9121  slate  pickers,  men  and  boys,  or  5.20  per  cent,  cleaning  10,046  tons 
per  employee  per  year,  or  41.51  tons  per  working  day. 

§  Based  on  the  full  tonnage;  namely,  91,626,964  tons. 


Tons  per 

Tons  per 

Employee 

Employees 

Per  Cent 

Employee 

per 

per  Year 

Working 

Day 

Pennsylvania  bituminous,  1913*: 

Miners       

122,932 

64.73 

1407 

5.61 

Other  inside  employees  

33,342 

17.56 

5188 

20.67 

Outside  employees  t 

33,635 

17  71 

5142 

20.49 

Total                   

189,909 

100.00 

911  1 

3.63 

*  All  mines  reporting. 

t  This  includes  10,122  coke  employees,  or  5.33  per  cent,  coking  37,381,029  tons,  or  3693 
tons  per  employee.  Without  these  men,  who  are  really  engaged  in  manufacturing,  there 
would  be  only  23  513  outside  employees,  or  12.38  per  cent,  the  production  being  7356  tons 
for  each  such  employee  per  year,  or  29.31  tons  per  working  day. 

J  Excluding  coke-workers,  this  figure  would  be  962. 


MINING  COSTS 


149 


TONNAGE  PER  EMPLOYEE  IN  PENNSYLVANIA,  WEST  VIRGINIA  AND  ILLINOIS 

Continued 


Tons  per 

Tons  per 

Employee 

Employees 

Per  Cent 

Employee 

per 

per  Year 

Working 

Day 

West  Virginia  report  of  1912  *: 

Miners  

43,581 

62.60 

1531 

6.84 

Other  inside  employees  

13,277 

19.07 

5026 

22.44 

Outside  employees  f  

12,758 

18.33 

5231 

23.35 

Total  

69,616 

100  .  00 

959 

4.28 

*  All  mines  reporting. 

t  This  includes  2297  coke  employees,  or  3.30  per  cent,  coking  3,310,250  tons,  or  1441  tons 
per  employee.  Without  these  men  there  would  be  only  10,461  outside  employees,  or  15.03 
per  cent,  the  production  being  6379  tons  per  year,  or  28.48  tons  per  working  day  for  each 
such  employee. 


Tons  per 

Tons  per 

Employee 

Employees 

Per  Cent 

Employee 

per 

per  Year 

Working 

Day 

Illinois  report  of  1913  *: 

Miners 

53588 

69  73 

1129 

6  31 

Other  inside  employees  

16,718 

21  75 

3619 

20  22 

Outside  employees 

6549 

8  52 

9240 

51  62 

Total  ...                   

76,855 

100  00 

787 

4  40 

*  Shipping  mines  only. 

under  this  caption  are  included  all  employees  inside  the  breaker, 
such  as  jigtenders,  platemen,  etc.  The  reports  from  the  bitumin- 
ous regions  of  Pennsylvania,  West  Virginia  and  Illinois  do  not 
give  any  enumeration  of  workmen  engaged  in  cleaning  and 
sizing,  though  there  are  many  men  so  employed,  some  being 
actually  on  picking  tables  and  others  trimmers  on  cars. 

Lest  too  much  weight  should  be  placed  on  the  cost   of 
breaker  work,  a  quotient  has  been  obtained  by  dividing  the 


150  COAL  MINING  COSTS 

whole  tonnage,  91,626,964  tons,  by  the  number  of  men  and 
boys  employed  on  the  outside,  excluding  those  working  within 
breakers  and  listed  as  slate  pickers.  The  divisor  is  thus  reduced 
from  46,643  employees  to  37,522.  The  quotient  to  which  we 
have  referred  is  2442  tons  per  year,  or  10.09  tons  per  day,  a 
figure  so  low  that  it  seems  hard  to  explain.  And  yet  in  com- 
paring this  figure  with  the  larger  figures  of  the  bituminous 
region,  it  must  be  borne  in  mind  that  the  latter  have  not  been 
"doctored"  by  the  omission  of  cleaners  and  sizers  in  making 
the  estimation. 

But  there  is  another  vitiating  principle  in  making  com- 
parisons to  which  attention  must  be  drawn.  In  the  anthracite 
region  there  are  many  strippings,  and  the  men  engaged  in  this 
work  would  find  their  appropriate  classification  in  our  tables, 
not  as  outside  men,  but  as  miners.  Unfortunately,  the  tonnage 
of  strippings  and  the  men  employed  at  such  open-cut  work 
have  not  been  segregated  in  the  mine  reports,  except  in  Illinois, 
and  there  the  stripping  work  is  of  almost  negligible  importance. 
In  the  year  ending  June  30,  1913,  about  152  men  at  seven  strip- 
pings in  Illinois  mined  137,448  tons. 

Another  difference  between  anthracite  and  bituminous  con- 
ditions makes  a  comparison  of  the  work  of  inside  men  some- 
what unfair.  In  the  anthracite  region  there  is  much  labor 
expended  in  reworking  old  breasts,  especially  in  the  Lansford 
district.  In  some  cases  possibly  there  are  advantages  in  this 
work,  making  it  more  profitable  than  first  mining,  but  in  nearly 
every  case  second  mining  of  this  description  is  far  more  expen- 
sive than  work  in  undisturbed  coal. 

Perhaps  there  are  no  sections  in  the  Union  with  more  effi- 
cient outside  handling  systems  than  those  found  in  western 
Pennsylvania  and  Illinois.  In  the  former  the  economy  largely 
arises  from  the  largeness  of  the  operations.  In  the  latter  it 
finds  its  source  possibly  in  the  fact  that  the  mines  are  mostly 
shafts  and  that  the  preparation  of  the  coal  is  limited  to  careful 
sizing.  The  fact  that  much  tramming  which  is  outside,  or 
tipple,  work  in  the  case  of  a  drift  is  inside,  or  caging,  work 
in  case  of  a  shaft,  also  accounts  for  the  lower  productivity 
of  Illinois  inside  hands.  The  gain  secured  on  the  surface  is 
a  loss  below. 

On  the  other  hand  in  western  and  central  Pennsylvania  the 


MINING  COSTS  151 

coal  instead  of  being  tipped  by  self-dumping  cages  is  dumped 
on  a  tipple,  which  takes  more  surface  hands  though  less  men 
are  needed  below.  The  long  haulages  common  to  drift  mines 
in  that  section  are  also  a  cause  of  wastage  of  labor,  though  they 
probably  more  than  pay  for  themselves  by  the  saving  in  shaft 
sinking  and  other  first  cost.  In  many  cases  there  are  long 
gravity  planes  which  need  the  services  of  many  men.  More- 
over, picking  tables  are  somewhat  numerous  in  central  Pennsyl- 
vania. 

The  necessity  that  shotfiring  in  Illinois  should  be  performed 
by  men  specially  employed  for  that  purpose  increases  the  num- 
ber of  men  engaged  in  underground  work.  In  many  cases  these 
men  are  supplemented  by  fire  hunters  whose  business  it  is  to 
see  that  the  shots  have  not  ignited  the  coal  and  to  extinguish 
an  incipient  blaze  if  such  be  found. 

It  will  be  noted  that,  excluding  coke  workers,  the  produc- 
tion per  outside  employee  in  the  bituminous  regions  of  Pennsyl- 
vania is  29.31  tons  per  day;  in  Illinois,  where  the  coke  ovens 
are  away  from  the  mines  and  are  therefore  not  considered  in 
the  state  report  of  the  mining  industry,  the  production  is  51.62 
tons  per  day  for  each  outside  workman. 

The  showing  for  this  latter  state  (Illinois)  is  truly  remark- 
able and  one  of  the  largest  operating  concerns  there  states  that 
it  is  not  obtained  by  excluding  men  in  official  position  from  the 
enumeration,  but  represents  a  real  economy  in  mine  labor. 
Irregular  operation  may,  and  frequently  does,  help  to  increase 
the  tonnage  of  coal  mined  per  miner  in  any  working  day,  but 
it  does  not  give  any  aid  to  the  management  in  increasing  effi- 
ciency in  other  labor;  for  when  places  are  vacated  by  the 
removal  of  workmen  who  become  discontented  at  the  frequent 
idlenesses,  the  transportation  problems  are  made  difficult  and 
the  outside  men  are  often  prevented  from  being  supplied  with 
the  tonnage  for  which  the  operation  was  designed. 

Some  sections  suffer  from  old-time  regulations.  In  the 
Blossburg  section  of  Pennsylvania,  for  instance,  it  was  and 
perhaps  still  is  customary  to  have  boys  at  all  switches,  though 
in  other  regions  the  mule-drivers  throw  their  own  switch  levers 
without  any  measurable  waste  of  time.  The  American  coal 
operator  concentrates  his  attention  on  this  question  of  tonnage 
per  man  employed  as  is  indicated  by  a  comparison  with  foreign 


152 


COAL  MINING  COSTS 


practice.     The  following  is  a  comparison  of  the  tonnage  per 
man  employed  in  the  various  fields  of  the  world: 


Pennsylvania  bituminous 1009 

Virginia 964 

West  Virginia 954 

Montana 893 

Wyoming 887 

New  Mexico 857 

Maryland 847 

U.  S.  bituminous 837 

Ohio 790 

Utah 783 

Illinois 775 

North  Dakota 773 

Indiana 772 

Colorado 770 

United  States 762 

Kentucky 745 

Alabama 720 

Washington 669 

Tennessee 613 

Australia 607 

New  Zealand 563 

Canada. .  529 


Pennsylvania  anthracite 520 

Transvaal 504 

Arkansas 480 

Iowa 478 

Texas 476 

Oklahoma 461 

Missouri 414 

Michigan 373 

Orange  Free  State 327 

German  Empire 301 

Natal 293 

Great  Britain 273 

Austria 231 

France 224 

Sweden 185 

Russia 173 

Belgium 173 

Spain 162 

Japan 133 

India 124 

Cape  of  Good  Hope 81 


The  figures  for  the  tonnage  per  man  in  the  United  States  are  taken  from 
the  report  of  the  U.  S.  Geological  Survey  for  1913;  those  for  all  other  countries 
are  for  1912,  except  in  the  case  of  Russia  (1909),  Spain  and  Japan  (1911). 
These  latter  figures,  of  course,  have  been  multiplied  by  1.12  so  as  to  obtain 
the  equivalent  in  short  tons. 

Of  course  this  order  of  precedence  in  economy  of  human 
labor  varies  from  year  to  year  according,  largely,  to  the  num- 
ber of  days  worked.  Thus,  for  instance,  in  1912  West  Virginia, 
being  second,  led  the  state  from  which  it  was  originally  dis- 
membered, and  Wyoming  stood  third  in  the  list.  In  1911  the 
output  per  man  in  Great  Britain  was  291  tons  and  that  of  the 
German  Empire  278  tons,  reversing  the  order  of  1912.  In 
many  states  and  countries  the  tonnages  are  low  because  of  the 
irregular  call  for  coal  where  the  main  demand  is  for  domestic 
use.  In  such  communities,  low  output  spells  not  inefficiency 
but  diversity  of  occupation.  The  farmer  has  added  mining  to 
his  ordinary  vocation. 

But  the  fact  that  every  state  in  this  Union  leads  every 
European  country  is  significant.  And  the  leadership  is  obtained 


MINING  COSTS  153 

despite  the  fact  that  the  mines  of  the  American  continent  do 
not  work  with,  by  any  means,  exemplary  regularity.  Still 
some  allowance  must  be  made  for  the  deeper  workings,  more 
contorted  measures,  more  perfect  preparation  of  coal  and  more 
back-filling  and  flushing  in  the  European  mines.  But  these 
difficulties  extenuate  rather  than  explain.  Smaller  cars  and 
haulage  units,  union  restrictions,  lower  speeds  and  less  use  of 
mining  machinery  are  the  real  reasons  for  the  lowered  human 
efficiency  in  the  coal  operations  across  the  Atlantic. 

In  fine,  it  has  been  observed  that  the  genius  of  the  American 
machine  is  in  capacity,  that  the  merit  of  the  English  is  in 
ruggedness  and  that  the  predominent  quality  of  the  German  is 
ingenuity  and  delicacy.  The  American  mechanism  performs 
great  service  with  little  watching.  The  English  machine  fre- 
quently does  less,  but  seems  to  live  forever.  The  German 
machine  is  often  complicated,  but  usually  not  needlessly  so. 
The  reason  for  the  complexity  is  that  the  German  has  a  delight 
in  modeling  those  types  of  machinery  which  demand  complica- 
tion and  refinement  of  detail. 

The  accompanying  chart,  Fig.  52,  excerpted  from  the  report 
of  the  Engineers  Committee  of  the  United  States  Fuel  Admin- 
istration gives  a  graphic  comparison  of  the  production  per  man 
in  the  United  States  and  Great  Britain  from  1896  to  1913. 
The  curves  for  both  anthracite  and  bituminous  are  given  for 
this  country  and  the  curves  of  the  number  of  men  employed 
both  inside  and  outside  are  given  for  Great  Britain. 

Miners'  wages. — It  may  be  argued  that  the  miners  are  mak- 
ing a  day's  pay  where  they  are  not  lazy  but  this  is  not  the 
case  in  many  mines.  The  operator  must  bring  pressure  to  bear 
upon  the  worker  to  get  him  to  work  regularly  and  do  a  full 
day's  work.  But  this  is  not  the  principal  difficulty,  and  .the 
difference  in  earnings  is  small  where  opportunities  to  load  coal 
are  reasonably  equal. 

The  rates  paid  for  mining  are  high — almost  too  high.  In  the 
union  districts  in  central  Pennsylvania  the  price  for  pick  coal 
(in  1915)  is  72c.  a  ton.  This  is  paid  whether  the  mining  is  in 
rooms  and  the  coal  hard  or  whether  it  is  in  pillars  which  have 
been  standing  and  the  coal  is  crushed.  In  the  latter  case  a 
man  can  dig  and  load  a  ton  in  20  min.  At  that  rate  he  ought 
to  earn  $8.64  a  day,  allowing  50  per  cent  of  the  time  for  rest- 


154 


COAL  MINING  COSTS 


>> 

,0 

ex 

I 
I 

a 
_o 

e 

1 

— 
*o 
c 


/ 


§ — — 


A! 


nr 


MINING  COSTS  155 

ing.  But  seldom  or  never  he  gets  enough  cars  to  earn  any- 
where near  that  amount. 

It  has  been  stated  that  the  men  in  central  Pennsylvania 
earned  more  money  when  pick  coal  was  35c.  a  ton  than  they  do 
now.  But  the  past  and  present  conditions  have  not  been  taken 
into  account.  The  blame  for  the  loss  in  earning  power  is  often 
angrily  put  on  the  union,  whereas  it  ought  to  be  partly  borne 
by  operators  for  not  maintaining  correct  conditions — men  in 
proportion  to  the  equipment  and  bosses  in  proportion  to  the 
mine. 

There  is  an  inflexibility  in  the  rates  paid  for  work  which 
makes  an  inequality  in  the  earning  power  of  the  men.  Rates 
are  generally  the  same  for  all  parts  of  a  mine  or  for  all  parts 
of  a  vein.  In  one  anthracite  mine  $1.58  is  paid  for  a  2-ton 
car  of  coal  whether  the  vein  is  8  ft.,  and  free  blowing  or  4  ft. 
and  hard  coal.  In  the  bituminous  regions  the  same  price  is 
paid  for  pick  coal  whether  it  is  in  rooms  or  in  crushed  pillars. 
Business  has  been  business  in  establishing  the  rates,  and  in 
any  argument  between  employer  and  employee  over  what  the 
rate  should  be,  the  attempt  on  both  sides  has  been  to  charge 
all  the  traffic  would  bear.  There  has  been  no  measurement  of 
work  by  which  attempts  could  be  made  scientifically  to  get  an 
equitable  basis  for  payment.  Measurement  of  work  is  just  as 
practical  in  mining  as  in  manufacturing,  where  it  is  now  being 
largely  done. 

Although  the  payment  is  made  according  to  the  ton  or  car 
loaded,  responsibility  for  the  amount  of  work  done  cannot  be 
wholly  cast  aside  by  the  management.  Men  are  used  in  a  mine 
entirely  too  freely  and  expected  to  be  on  their  own  resources. 
They  cannot  do  it,  for  they  do  not  work  alone,  but  are  depend- 
ent upon  others  for  the  opportunity  to  work. 

The  drivers,  being  paid  by  the  day,  are  not  anxious  to 
deliver  more  cars  than  are  necessary  to  keep  their  jobs  and 
so  neglect  the  miner.  In  consequence  the  miner  does  .not  earn 
as  much  as  he  might.  One  can  hardly  blame  the  driver  so 
long  as  it  is  a  fixed  principle  in  life  to  get  as  much  for  as  little 
as  possible.  Universally  a  mine  is  so  undermanned  by  bosses 
that  no  work  can  be  followed  up,  and  the  neglect  of  duty  and 
failure  to  cooperate  can  become  so  common  that  the  miners 
frequently  suffer  from  needless  delays. 


156  COAL  MINING  COSTS 

It  is  hard  to  stir  up  trouble  at  a  mine  where  the  men  are 
being  properly  treated  and  are  earning  a  day's  pay.  It  will 
be  found  that  the  strength  of  a  union  increases  at  a  mine  in 
proportion  as  there  is  more  cause  for  discontent.  But  this  is 
not  to  be  regarded  as  an  argument  that  the  men  should  be 
paid  whether  they  do  any  work  or  not. 

Increased  earnings  of  the  individual  can  be  reflected  favor- 
ably on  the  cost  sheet.  A  man  who  is  working  a  piece  of  coal 
which  has  been  formerly  neglected  and  is  so  situated  that  he 
is  given  only  a  car  or  two  a  day  and  whose  earnings  in  con- 
sequence are  low  is  likely  to  be  a  source  of  trouble.  When  it 
is  necessary  to  keep  a  man  and  a  mule  for  only  a  few  cars 
a  day,  the  transportation  charge  will  be  high.  The  cost  of 
gathering  was  taken  in  one  mine  and  found  to  run  from  2y2c. 
to  25c.  The  total  cost  for  haulage  was  high  because  on  only 
a  few  roads  were  there  enough  men  to  give  a  driver  a  full  day 's 
work.  When  the  work  was  apportioned  so  that  there  were 
enough  men  on  each  road,  the  cost  fell. 

It  would  be  more  advantageous  to  the  workmen  if  the  union 
leaders,  instead  of  making  demands  for  further  increase  in  the 
rates  of  pay,  asked  for  a  better  organization  of  the  work  so 
that  the  men  were  furnished  an  equal  number  of  cars.  Rates 
are  high  enough  so  that  no  man  need  overwork  himself  in 
earning  good  pay.  In  a  boiler-room  a  fireman  will  shovel  into 
the  furnace  from  18  to  20  tons  a  day,  which  is  more  than  a 
miner  will  load  into  cars. 

One  well-known  union  leader  has  been  quoted  as  saying 
that  15  tons  was  not  too  much  for  a  man  loading  after  machines, 
and  the  union  representatives  agreed  to  a  task  of  12  tons  at 
one  mine.  When  one  gets  into  the  complex  changing  condi- 
tions of  an  anthracite  mine  it  might  seem  a  little  doubtful 
whether  this  could  be  done,  but  again  it  is  a  case  of  the  arrange- 
ment of  work  rather  than  a  crying  need  of  an  increase  in 
rates. 

Under  the  present  general  manner  of  payment  the  only 
cure  is  to  promulgate  the  idea  that  each  man  should  produce 
a  large  tonnage  per  day.  An  increase  in  their  output  will  be 
of  advantage  to  the  miners,  as  they  are  paid  by  the  ton.  It 
will  also  be  to  the  interest  of  the  company,  for  an  increased 
output  per  handling  unit  will  result  from  the  proper  organiza- 


MINING  COSTS  157 

tion  of  work  by  which  the  larger  product  is  secured.  Though 
there  is  no  one  cure  for  all  ills,  the  application  of  the  principles 
of  efficiency  to  mining  will  greatly  help  to  alleviate  the  con- 
ditions which  are  now  a  frequent  and  natural  source  of  much 
irritation. 

The  accompanying  table  gives  the  wage  scale  in  effect  in 
the  Hocking  District  from  1898  to  1921,  covering  all  classes  of 
labor  in  and  about  the  mines.  This  table  will  be  found  of  use 
in  computing  the  comparative  cost  of  doing  the  various  kinds 
of  work  cited  throughout  this  book  where  the  figures  are  not 
up  to  date. 

The  Hocking  scale  as  it  is  generally  known  has  been  used  for 
this  purpose  for  the  reason  that  it  dates  back  the  farthest 
of  any  of  the  important  scales  and  also  because  it  is  used  as 
a  base  for  all  wages  in  the  Central  competitive  field  and  has 
influenced  wage  settlements  in  all  union  fields  generally.  The 
break  in  the  table  in  1914  was  occasioned  by  the  change  from 
the  screened  coal  basis  of  payment  to  the  mine  run. 

Losses  from  idle  time. — The  yearly  average  of  days  of 
activity  in  the  mines  of  the  United  States  from  1908  to  1914 
was  217.  Thus  the  mining  plants  averaged  only  about  72  per 
cent  of  full  time.  In  the  year  1914  the  number  of  working 
days,  according  to  United  States  Geological  Survey  report,  was 
only  207  for  all  coal  mines  throughout  the  country. 

Unfortunately  even  this  time  is  not  by  any  means  evenly 
distributed.  Some  mines  are  idle  a  large  percentage  of  the 
year,  while  others  work  with  regularity.  A  few  mines  in  agri- 
cultural states  work  mostly  in  the  winter,  and  in  the  summer 
the  men  must  find  farming  jobs  to  keep  them  busy,  though  it 
is  doubtful  whether  many  of  them  do.  A  miner  is  not  always 
willing  to  work  under  a  hot  sun.  Arkansas  and  Oklahoma  in 
1913,  for  instance,  only  worked  174  days,  or  barely  55.5  per 
cent  of  the  possible  working  time. 

Perhaps  under  no  circumstances  will  it  be  possible  to  keep 
coal  mining  at  an  even  pace  throughout  the  year,  but  it  should 
be  possible  to  do  better  than  the  1914  season.  Such  depression 
as  occurred  in  that  year  would  be  helped  by  a  combination 
embracing  all  the  mine  operators,  but  the  conditions  of  unem- 
ployment were  nation-wide,  and  something  broader  than  a  com- 
bination of  mine  owners  would  be  necessary  as  a  stabilizer. 


158 


COAL  MINING  COSTS 


WAGE  SCALES  IN  THE  HOCKING  DISTRICT,  CENTRAL 


1892 
to 
1894 

Feb. 
to 
April, 
1894 

1894 
to 
1895 

June 
to 
Oct., 

1895 

1895 
to 
1896 

March 
to 
Oct., 
1896 

Pick  Mining 
Screened  lump,  per  ton  
Mine  run,  5-7  lump  price  
Entry,  <lry,  per  yard  

$0.70 
.50 
1  75 

$0.50 
.35* 
1  25 

$0.60 
.42f 
1  50 

$0.51 
.36| 
1  274 

$0.55 
.39f 
1  37$ 

$0.61 
.43* 
1   52$ 

Entry  —  breakthroughs,  yard.  .  .  :  

1  75 

1  25 

1  50 

1  274 

1  37$ 

1   52$ 

Room  —  breakthroughs,  yard  
Room  turning,  per  room  

1.00 
2.50 

.75 
2  50 

1.00 
2  50 

.77$ 
2  50 

.87$ 
2  50 

1.02$ 
2  50 

Track  layers,  per  day  

2.25 

1  75 

2  00 

1  78$ 

1  87$ 

2  02$ 

Track  layers'  helpers,  per  day  

Trappers,  per  day  

.75 

75 

75 

75 

75 

75 

2  00 

1  50 

1  75 

1  52$ 

1  62$ 

1   774 

Trip  (rope)  riders,  per  day  

1  75 

1  50 

1  75 

1  52$ 

1  62$ 

1  75 

Timber  men,  per  day  

Pipe  men,  per  day  

Wire  men,  per  day  

Motor  men,  per  day  

Pumpers   per  month 

40  00 

30  00 

35  00 

30  00 

32  50 

35  00 

Other  inside  day  labor  

Outside 
First  blacksmiths,  per  day.    .  .  . 

Second  blacksmiths,  per  day 

Blacksmiths'  helpers,  per  day 

x 

Carpenters,  per  day 

x 

Dumpers-trimmers,  per  day   .  .  . 

2  00 

1     0 

1  75 

1  52$ 

1     2$ 

1  77* 

Slack  haulers,  per  day  .    .    . 

Greasers,  couplers,  per  day.  .  . 

x 

Engineers,  firemen,  per  day. 

x 

Other  outside  day  labor    .  .  . 

1  75 

1  25 

1  50 

1  25 

1  37$ 

1  50 

Machine 
Cutting,  per  ton  —  by 
Jeffrey  styles  —  room  

08 

07 

08 

07 

071 

08 

Jeffrey  styles  —  entry  

11 

10 

11 

10 

10 

11 

Punch  machines  —  room  

12$ 

Hi 

12$ 

11$ 

12 

124 

Punch  machines  —  entry  

.13$ 

12$ 

13$ 

12$ 

13 

134 

Loading,  per  ton  — 
In  rooms  

.35 

25 

30 

25$ 

27$ 

30$ 

In  rooms  with  hand  drilling  

.38 

28 

33 

28$ 

30$ 

334 

In  entry  

.43$ 

Sli1* 

36 

31$ 

33$ 

364 

In  entry  with  hand  drilling  

.46$ 

34A 

39 

34$ 

36$ 

394 

In  breakthroughs  in  entry.  .  .  .Entry.  .  .  . 

Pr^ce 
41 

E.  P. 
29f 

E.  P. 
35 

E.  P. 
29$ 

E.  P. 

32^r 

E.  P. 

35A 

In  breakthroughs  in  rooms    with     hand 
drilling 

41 

32f 

38 

32$ 

35^nr 

38A 

Drilling,  by  hand,  per  ton    . 

03 

03 

03 

03 

03 

03 

by  machine,  per  ton 

02 

014 

02 

01$ 

01  3 

02 

Machine  by  the  day  Runner  and 
helper  jointly 

4  50 

4  50 

Room  turning  to  cutter  and  loader.     Entry 
price  

E.  P. 

E.  P. 

E.  P. 

E.  P. 

*  Formerly  rated  with  couplers,  greasers,  etc. 


MINING  COSTS 


159 


COMPETITIVE  YIELD,  FOR  THE  YEARS  1892  TO  1921 


1896 
to 
1897 

Jan. 
to 
July, 

1898 

1897 

to 
1898 

1898 
to 
1900 

1900 
to 
1903 

1903 
to 
1904 

1904 
to 
1906 

1906 
to 
1908 

1910 
to 
1912 

1912 
to 
1914 

$0.45 

$0.51 

$0.56 

$0.66 

$0.80 

$0.90 

$0.85 

$0.90 

$0.95 

$1.00 

.32$ 

.36f 

.40 

.47* 

.57* 

.64f 

.60f 

.64| 

.6785 

0.7lf 

1.124 

1.274 

1.40 

1.65 

2.00 

2.25 

2.124 

2.25 

2.3748 

2  .  4996 

1.124 

1.274 

1.40 

1.65 

2.00 

2.25 

2.124 

2.25 

2  .  3748 

2  .  4996 

.624 

.774 

.90 

1.15 

1.39 

1.56 

1.474 

1.56 

1  .  6465 

1  .  7330 

2.50 

2.50 

2.50 

2.50 

3.03 

3.41 

3.22 

3.41 

3  .  5992 

3.7884 

1.65 

1.784 

1.90 

1.90 

2.28 

2.56 

2.42 

2.56 

2.70 

2.84 

1.75 

2.10 

2.36 

2.23 

2.36 

2.49 

2.62 

.75 

.75 

.75 

.75 

1.00 

1.13 

1.064 

1.13 

1.25 

1.32 

1.40 

1.524 

1.65 

1.75 

2.10 

2.56 

2.42 

2.56 

2.70 

2.84 

2.56 

2.42 

2.56 

2.70 

2.84 

1.40 

1.524 

1.65 

1.75 

2.10 

2.56 

2.42 

2.56 

2.70 

2.84 

* 

* 

* 

* 

2.10 

2.56 

2.42 

2.56 

2.70 

2.84 

1.90 

2.28 

2.56 

2.42 

2.56 

2.70 

2.84 

1.85 

2.22 

2.50 

2.36 

2.50 

2.63 

2.78 

2.56 

2.42 

2.56 

2.70 

2.84 

2.56 

2.42 

2.56 

2.70 

2.84 

30.00 

30.00 

33.00 

o 

2.10 

2.36 

2.23 

2.36       2 

2.49 

2.62 

CO 

2.81 

2.654 

2.81       .g 

2.96 

3.12 

2.53 

2.39 

2.53        S 

2.67 

2.81 

2.36 

2.23 

2.36       S 

2.49 

2.62 

2.53 

2.39 

2.53        ° 

2.67 

2.81 

1  40 

1  524 

1  65 

1.75 

2.10 

2.36 

2.23 

2.36       oo 

2.49 

2.62 

1.97 

1.86 

1.97       8 

2.07 

2.18 

1.41 

1.33 

1.41       1 

1.48 

1.56 

124%  } 

Reduced  4 

Adv.      ~ 

Adv. 

Adv. 

1.25 

1.25 

1.40 

* 

* 

adv.    / 

adv.  of  1903 

2.36       '§, 

2.49 

Adv. 

1 

.06 

.07 

.074 

.08 

.09 

.10 

.09 

.0965  | 

.1030 

.1050 

.09 

.10 

.104 

.11 

.121 

.134 

.12f 

.1320  W 

.14 

14.35 

.104 

.114 

.12 

.124 

.134 

.144 

.14 

.144    1 

.15 

15.44 

.114 

.124 

.13 

.134 

.141 

.16 

.151 

•16        | 

.16| 

.1710 

.224 

.254 

.28 

.33 

.41 

.48 

.45 

.4835  tf 

.5170 

.5550 

.254 

.28j 

.31 

.36 

.44 

.51 

.48 

.5135 

.5470 

.5850 

.284 

.31j 

.34 

.414 

.514 

.60 

.56i 

.6030 

.6435 

.6885 

.314 

.344 

.37 

.444 

.544 

.63 

.59i 

.6330 

.6735 

.7185 

E  P 

E.  P. 

E.  P. 

E.  P 

E.  P 

E   P 

E.  P. 

E.  P. 

E.  P. 

E.  P. 

.264 

.294 

.324 

139 

.481 

.56-^0 

.52^ 

.5650 

.6030 

.6455 

.294 

.324 

.354 

.42 

.511 

.59tin> 

.5511, 

.5950 

.6330 

.6755 

.03 

.03 

.03 

.03 

.03 

.03 

.03 

.03 

.03 

.03 

.014 

014 

.01} 

02 

.021 

.024 

021 

02* 

.0263 

.0276 

4.20 

4.50 

.  V/A'g 

•  v**  j 

E.  P. 

E.  P. 

E.  P. 

E.  P. 

E.  P. 

E.  P. 

E.  P. 

E.  P. 

E.  P. 

E.  P. 

x  Special  prices  according  to  nature  of  work. 


160 


COAL  MINING  COSTS 


WAGE  SCALES  IN  THE  HOCKING  DISTRICT  FOR  THE  YEARS  1892  TO  1921 — Continued 


1914 
to 
1916 

1916 
to 
1918 

April  to 
Oct., 
1917 

1917 
to 
1920 

1920 
to 
1921 

Pick  Mining 
Run  of  mine,  per  ton  

$0.676 
2.4996 
2.4996 
1  .  7330 

3.7884 

2.84 
2.62 
1.32 
1.50 

2.84 
2.84 
2.84 
2.84 
2.78 
2.84 
2.84 

$0.6764 
2.6245 
2.6245 
1.8196 
3.978 

2.98 
2.75 
1.40 
1.59 
2.98 
2.98 
2.98 
2.98 
2.92 
2.98 
2.98 
2.98 
2.75 

.074 
.  10235 
.1084 
.  12016 

.406 
.426 
.49945 
.51945 
.51945 
.489 

.02 
.0195 

3.27 
2.95 
2.75 
2.95 
2.75 
2.28 
1.64 

$0.7764 
2  .  6245 
2.6245 
1.8196 
3.978 

3.60 
3.35 
1.90 
2.19 
3.60 
3.60 
3.60 
3.60 
3.52 
3.60 
3.60 
3.60 
3.35 

.0890 
.1173 

$0.8764 
3.0181 
3.0181 
2.0925 
4.5747 

5.00 
4.75 
2.65 
3.59 
5.00 
5.00 
5.00 
5.00 
4.92 
5.00 
5.00 
5.00 
4.75 

.1040 
.1365 
.1384 
.1518 

.5760 
.5960 
.6835 
.7035 
.7035 
.6684 

.02 
.0195 

5.27 

4.95 
4.75 
4.95 
4.75 

3.24 

1    3 

R  >>  i  «# 

3111* 
qiiii 

$1.1164 
3.6217 
3.6217 
2.5110 
5.4896 

6.00 
5.75 
3.18 
4.59 
6.00 
6.00 
6.00 
6.00 
5.92 
6.00 
6.00 
6.00 
5.75 

.14 
.1790 
.1744 
.1905 

.80 

.9290 
.9290 
.9290 

6.27 
5.95 
5.75 
5.95 
5.75 

4.24 

Entries,  dry,  per  yard  

Inside  Day  Labor 

(Where  old  men  are  employed) 

Bottom  cagers,  drivers,  trip  riders,  per  day.  . 
Snappers  on  gathering  locomotives,  per  day.  . 
Water  haulers,  machine  haulers,  per  day.  .  .  . 
Timber  men,  per  day  

Pipe  men,  for  compressed  air  plants,  per  day. 

All  other  inside  day  labor,  per  day  
Spike  team  drivers,  25  cents  per  day  extra.  .  . 

Machine  Cutting 
By  Jeffrey  style  of  machines,  in  room,  per  ton. 
By  Jeffrey  style  of  machines,  in  entry,  per  ton. 
By  punching  machines,  in  rooms,  per  ton  
By  punching  machines,  in  entries,  per  ton.  .  . 

Loading 

2.62 

.07 
.0970 
.1044 
.1156 

.38 
.40 
.4690 
.4890 
E.  P. 
.46 
.44 

.02 

.0186 
E.  P. 

3.12 
2.81 
2.62 
2.81 
2.62 
2.18 
1.56 

.4910 
.5110 
.5845 
.6045 
.6045 
.5740 

In  rooms  with  hand  drilling,  per  ton. 

In  entry  with  hand  drilling,  per  ton 

Breakthroughs  in  entry,  per  ton 

Breakthroughs  in  rooms  

Drilling 
By  hand  per  ton  

By  machine,  per  ton   

3.87 
3.55 
3.55 
3.55 
3.35 
2.88 
2.24 

^IJ    - 

;N 

i^3 

Room  turning,  cutter  and  loader 

Outside  Day  Wage  Scale 
First  blacksmith,  per  day  

Second  blacksmith,  per  day  

Blacksmith  helpers,  per  day  

Carpenters,  per  day  

Dumpers  and  trimmers,  per  day  

Slack  haulers,  per  day  

Greasers  and  couplers,  per  day  

Where  engineers  and  firemen  are  employed 
by  the  day,  the  minimum  rate  is  $4.75  for 
8  hrs.     This  does  not  apply  to  men  employed 
at  a  monthly  rate.     This  also  applies  to  coal 
washers. 

MINING  COSTS 


161 


The  whole  nation  must  get  together  to  produce  a  stability  in 
business  which  will  make  steady  work  in  coal  mines  and  in 
every  other  form  of  activity. 

The  case  of  the  miner  against  irregular  operation  has 
already  been  forcibly  set  before  the  public.  What  is  not  so 
generally  realized  is  that  the  case  of  the  operator  is  just  as 
damaging  to  him.  His  capital  is  idle  and  his  mine  equipment, 
instead  of  benefiting  by  a  rest,  is  rapidly  depreciating.  Al- 


1.0  U 

1.50 
1.40 
UO 

1  on 

| 

/ 

/ 

/ 

Percentage  Increase  in  Cosf 
3  S  s!  2  S  g  5  ; 

/ 

/ 

/ 

/ 

/ 

/ 

/ 

/ 

/ 

> 

.40 
.30 
.W 

/I 

x 

X 

^ 

^ 

.10 
\ 

^  —  ' 

^ 

^-'l 

^ 

H-O^ 

r**^ 

0    95     90     85     80     15      TO     65     60      55     50     45     4        35     30    ZS     Z       IS      1C 

5      0 

Percen-fags  Decre&ecil  Car  Supply 

FIG.  53. — Percentage  of  increase  cost  due  to  irregular  working  time  compiled 
from  830  observations  in  the  New  River  field. 

though  the  mine  shuts  down,  his  fixed  charges  run  on — not 
only  interest  charges  and  salaries,  but  a  host  of  maintenance 
charges  as  well.  And  in  the  end  the  coal  consumer  pays  the 
bill  for  idleness  of  miner  and  mine. 

In  this  connection  we  may  find  instruction  in  an  exceed- 
ingly valuable  study  made  by  Messrs.  Garnsey,  Allport,  and 
Norris  of  the  costs  of  production  as  effected  by  interruptions 
of  working  time.  Fig.  53  is  taken  from  their  "  Report  of  the 
Engineers'  Committee  of  the  United  States  Fuel  Administra- 


162  COAL  MINING  COSTS 

tion,  1918-19. ' '  It  represents  an  analysis  of  the  monthly  records 
of  seventy-three  operators  in  the  New  River  district  of  West 
Virginia.  Each  of  these  was  carefully  analyzed,  and  the  per- 
centage increase  of  cost  for  each  of  the  830  observations  thus 
obtained  was  plotted;  weighed  averages  were  then  taken  at 
each  2.5  per  cent  from  70  to  100  per  cent  working  time,  and 
for  each  5  per  cent  below  80  per  cent.  The  result  of  this 
study  is  shown  in  Fig.  53,  which  has  been  checked  by  numerous 
observations  from  practically  every  field  and  has  been  found, 
within  reasonable  limits,  to  be  correct.  This  diagram  can  and 
has  been  used  in  reducing  to  normal  cost  the  reported  cost  of 
collieries  shut  down  during  parts  of  months.  The  reason  for 
the  increased  cost  per  unit  of  output  is,  that  the  smaller  the 
number  of  tons  produced  the  larger  is  the  share  of  fixed  over- 
head expenses  that  must  be  borne  by  each  ton.  F.  S.  Peabody 
testifying  before  the  Frelinghuysen  committee  in  1919  stated 
that:  "the  earning  of  the  laborer  and  the  cost  of  coal  depend 
entirely  on  continuous  work.  Our  costs  will  vary  from  month 
to  month,  dependent  on  the  running  time  of  our  mines.  There 
will  be  a  variation  of  between  50  and  60c.  a  ton  from  month 
to  month,  depending  on  the  number  of  hours  the  mines  are 
idle." 

The  average  number  of  productive  days  worked  per  annum 
in  Illinois  and  Indiana  is  only  about  175  out  of  a  possible  300 
or  more.  This  idle  time  of  the  miners  is  not  confined  to  one 
season  or  period  during  which  they  can  find  employment  else- 
where. The  men  are  always  subject  to  call,  for  which  reason 
they  urge  a  greater  daily  wage  so  that  that  their  annual  in- 
come may  be  sufficient  for  their  needs.  This  causes  these 
operators  to  grant  abnormal  wage  advances,  which  are  directly 
reflected  in  coal  cost. 

Many  industrial  plants  which  produce  standard  or  basic 
commodities  find  it  possible  to  operate  24  hr.  per  day  by  using 
different  shifts  of  men.  They  work  also  for  310  or  more  days 
a  year,  or  a  total  of  7440  hr.  per  annum.  Still  other  industries, 
on  two  8-  or  10-hr,  shifts  per  24  hr.  work  300  to  310  days  per 
year,  thus  operating  5000  to  6000  hr.  every  12  months. 

Even  one  8-hr,  shift  in  each  24-hr,  period  with  310  days 
per  year  gives  2480  working  hours  in  every  12  months.  Be- 
cause of  the  unrestricted  competition  the  mine  operators  of 


MINING  COSTS  163 

Illinois  and  Indiana  have  built  more  plants  than  are  needed 
and  can  only  operate  for  8-hr,  out  of  every  24,  and  for  175 
days  per  year,  or  1400  hr. 

It  will  be  seen,  therefore,  that  as  against  100-per  cent  plant 
utilization  (24  hr.  for  310  days  or  7440  hr.  per  annum)  pos- 
sible in  some  industries,  and  as  against  an  average  by  all 
industries  of  33  per  cent  to  45  per  cent  (one  8  or  10-hr,  shift 
per  24-hr,  period  for  310  days),  a  coal  plant  is  in  actual  pro- 
ductive use  only  about  18  per  cent  of  the  mine.  This  makes 
plant,  interest  and  depreciation  charges  six  times  as  heavy  as 
for  some  other  industries. 

The  97,000  miners  of  Illinois  and  Indiana  who  are  prevented 
from  working  125  days  per  year  might  at  the  1915  wage  have 
earned  an  additional  $36,400,000,  or  $371  per  man  per  year, 
had  their  employers  been  able  to  give  them  work  or  had  their 
efforts  been  expended  in  other  directions. 

Because  the  operators  can  give  their  miners  work  only  a 
part  of  the  time,  these  operators  must  pay  higher  daily  wages 
than  are  warranted  by  the  current  selling  prices.  Their  labor 
cost  (in  1915)  is  92.44c.  per  ton,  whereas  the  selling  price  is 
but  $1.14  and  $1.11  respectively  for  the  states  of  Illinois  and 
Indiana. 

Statistics  of  the  coal-mining  industry  for  1912,  give  a  total 
production  of  bituminous  coal  of  450,000,000  tons,  with  a  total 
production  cost  per  ton  as  follows : 

Direct  labor  per  ton $0 . 5425 

Indirect  labor  (day  work) 0. 2383 

Salaries 0. 0575 

Supplies 0. 1305 

Royalties 0.0320 

Miscellaneous  expenses 0. 0530 

Total  per  ton $1.0538 

Assuming  that  as  claimed  by  competent  authorities,  the 
output  from  the  same  workings  might  have  been  600,000,000 
tons,  an  increase  of  33V3  per  cent,  with  little  advance  in  the 
expenditures,  let  us  see  how  this  increase  in  output  would 
affect  the  production  cost  per  ton. 

Direct  labor  and  royalties  probably  would  remain  about 


164  COAL  MINING  COSTS 

as  before,  but  the  other  expense  items  would  show  a  reduction 
due  to  the  increased  output  about  as  follows : 

Direct  labor  per  ton $0 . 5425 

Indirect  labor  (day  work) 0. 1790 

Salaries 0.0432 

Supplies. 0. 0978 

Royalties 0. 0320 

Miscellaneous  expenses 0 . 0400 


Total  per  ton $0.9345 

This  shows  a  decrease  in  production  cost  per  ton  of  11.93c., 
or  11.4  per  cent,  due  simply  to  increase  of  tonnage  produced. 

Economic  aspects  of  conservation. — It  has  been  customary 
to  refer  to  the  unlimited  coal  resources  of  the  United  States, 
but  while  our  country  has  been  wonderfully  blessed  in  this 
respect,  the  exhaustion  of  some  of  the  choicest  coal  beds  is 
already  in  sight,  as  for  instance,  the  high-grade  coking  coals 
of  the  Connellsville  region.  Conservative  estimators  realize 
that  the  coal  supply  should  now  probably  be  measured  by 
hundreds  rather  than  by  thousands  of  years,  and  that  it  be- 
hooves us  to  conserve  our  fuel  resources.  Large  areas  of  high- 
grade  fuel  undoubtedly  yet  remain,  and  it  is  not  too  late  to 
utilize  these  deposits  much  more  fully  than  has  been  done 
with  similar  deposits  in  the  past,  both  by  more  skillful  mining 
and  by  a  more  thorough  utilization  of  the  coal  after  it  is 
brought  to  the  surface. 

The  Anthracite  Coal  Waste  Commission  in  1893  estimated 
that  probably  not  over  30  to  35  per  cent  of  the  coal  originally 
contained  in  the  areas  mined  over  has  been  saved,  a.nd  that 
even  by  working  over  the  old  culm  banks  and  reworking  the 
area  already  mined,  not  over  10  per  cent  additional  would  be 
obtained,  thus  giving  a  loss  of  50  to  60  per  cent  of  the  original 
coal.  By  means  of  the  very  close  washing  and  sorting  of  the 
small  sizes,  and  the  better  removal  of  the  pillars  through  the 
crushing  of  culm  and  other  methods,  the  amount  of  waste  may 
be  somewhat  less  than  is  estimated  by  the  Coal  Waste  Com- 
mission, but  still  an  enormous  waste  is  going  on,  of  a  material 
which  is  not  duplicated  anywhere  else  in  the  country,  and  of 
which  the  supply  is  comparatively  limited. 

Dr.  I.  C.  White,  has  estimated  about  1909,  that  in  mining 


MINING  COSTS  165 

bituminous  coal  in  the  United  States  not  over  50  per  cent  of 
the  coal  in  the  ground  has  been  obtained.  This  figure  is  exces- 
sive for  present-day  practice,  still  the  amount  of  waste  is 
entirely  too  great  and  should  be  decreased. 

That  great  portions  of  our  coal  deposits  have  been  skimmed 
over,  leaving  rich  territory  abandoned  because  of  poor  and 
inadequate  methods,  which  aimed  at  the  recovery  of  the  easily 
accessible  and  abandoned  the  more  difficult,  is  apparent  to 
every  observer.  Much  valuable  coal  property  has  thus  been 
wasted  or  ruined. 

It  is  in  a  sense  a  moral  obligation  on  the  operator  to  recover 
the  pillar  and  top  coal  that  the  loss  to  the  country  may  be 
lessened,  but  where  this  involves  an  additional  expense,  it  can- 
not be  undertaken.  Still,  for  every  two  acres  of  coal  land 
which  the  operations  exhaust  they  leave,  one  acre  of  coal 
unrecovered  and  unrecoverable  in  the  ground.  This  means  that 
in  Illinois  each  year  12,000  acres  of  coal  land  are  exhausted, 
whereas  the  exhaustion  should  be  but  8000  acres.  In  Indiana 
the  depletion  is  3000  acres,  whereas  it  should  not  be  more  than 
2000.  In  the  whole  country  100,000  acres  are  exhausted, 
whereas  not  more  than  65,000  or  70,000  acres  should  have  been 
thus  made  of  no  mineral  value. 

A  summary  of  the  recovery  effected  in  the  different  mining 
fields  will  give  a  basis  for  fixing  a  standard  recovery  to  work 
towards.  Conditions  prevailing  in  most  of  the  important  min- 
ing fields  in  1914  were  described  in  a  paper  presented  before 
the  West  Virginia  Mining  Institute  in  that  year. 

In  order  to  get  an  idea  of  what  is  being  accomplished  in 
other  fields,  inquiries  were  sent  out  to  different  sections  of  the 
United  States.  The  results  are  shown  in  a  condensed  form  by 
the  accompanying  table. 

It  will  be  noted  on  this  table,  the  wide  variation  of  per- 
centages given  for  different  districts  in  various  states.  All  are 
large  producers  of  coal,  with  one  or  two  exceptions,  and  employ 
what  are  presumed  to  be  modern  methods  of  mining.  It  is 
noticeable  that  the  thin  seams  usually  are  overlaid  with  good 
roof  and  the  percentage  of  recovery  is  high.  Also,  that  but 
one  operator  expects  the  ultimate  recovery  to  fall  below  his 
present  percentage. 

In  the  southern  Colorado  field,  where  the  roof  and  bottom 


166  COAL  MINING  COSTS 

conditions  are  favorable  for  pillar  drawing,  no  roof  coal  is  left 
for  protection  and  the  recovery  is  given  as  80  to  90  per  cent, 
working  on  the  room-and-pillar  system.  It  is  claimed  that  in 
the  Canon  City  district,  where  the  longwall  system  is  used, 
that  100  per  cent  of  the  seam  is  recovered.  This  is  in  the 
thinner  seam,  which  measures  about  3  ft. 

Rooms  in  the  southern  Colorado  district  are  driven  16  to 
18  ft.  wide  on  45-ft.  centers,  while  in  the  Walsenburg  district, 
where  the  coal  is  harder  and  has  less  cover,  rooms  are  driven 
35  to  40  ft.  wide,  leaving  the  same  thickness  in  pillars  which 
are  recovered  by  machine  and  pick  work.  Track  is  laid  on  each 
side  of  the  room  and  frequently  one  or  two  cuts  are  taken  off 
the  side  of  the  pillar  with  a  machine  before  beginning  pick 
work. 

The  bottom  of  the  Colorado  seams  is  usually  slate  of  a  soft 
character,  which  heaves  when  weight  is  thrown  onto  the  pillars, 
making  it  necessary  frequently  to  drive  a  skip  along  the  pillar 
in  order  to  reach  the  back  end  before  beginning  to  draw  it. 

There  are  districts  where  both  roof  and  bottom  conditions 
are  unfavorable  and  much  difficulty  is  encountered  in  breaking 
the  overlying  strata.  In  these  sections  the  recovery  is  esti- 
mated to  be  60  to  65  per  cent ;  15  or  20  per  cent  is  lost  in  roof 
coal  because  the  strata  next  overlying  the  coal  cannot  be 
propped. 

Another  company,  operating  in  practically  the  same  field, 
states  its  recovery  runs  75  to  80  per  cent  of  the  entire  seam, 
while  a  higher  ultimate  recovery  is  expected.  This  firm  is  now 
driving  room  entries  to  the  boundaries  and  the  last  rooms  are 
worked  first,  thus  making  it  possible  to  draw  the  pillars  on 
the  retreat. 

In  the  Michigan  field,  especially  in  the  Saginaw  district, 
the  coal  is  in  pockets  rather  than  a  continuous  seam.  The 
basin  lies  for  the  most  part  in  a  low,  flat  country,  and  shafts 
about  200  ft.  deep  are  necessary  to  reach  the  coal.  The  bed 
averages  about  3  ft.,  and  is  of  poorer  grade  than  the  Ohio 
and  Pennsylvania  fuels,  so  that  its  market  is  somewhat  limited. 

The  top  in  these  mines  is  usually  black  slate,  while  one 
mine  has  a  fireclay  roof,  making  it  necessary  to  leave  top  coal. 
Yet  rooms  are  driven  40  ft.  wide  with  track  along  each  rib. 
The  length  of  the  room  is  150  ft.,  as  the  miner  pushes  his  cars 


MINING  COSTS  167 

from  the  working  place  to  the  entry.  With  the  conditions 
just  given,  the  recovery  claimed  is  between  80  and  90  per  cent. 
The  65  per  cent  recovery  given  in  the  table  (Item  4)  repre- 
sents the  result  of  leaving  pillars  for  surface  protection  within 
the  city  limits. 

Going  into  central  Illinois  fields,  where  the  No.  6  seam  is 
operated  extensively,  there  are  adverse  public  feelings  and 
unsettled  industrial  and  labor  conditions,  which  materially 
affect  the  percentage  of  recovery.  Surface  costing  $100  to 
$250  per  acre  cost  the  operator  two  or  three  times  these  values 
in  cases  of  subsidences,  if  the  mining  rights  do  not  clearly 
cover  the  property.  Besides  these  factors,  the  companies 
operating  in  the  Glen  Carbon,  Mt.  Olive  and  Divernon  fields, 
state  that  owing  to  thick,  soft  clay  under  the  coal  or  great 
overburden  (300  ft  to  400  ft.)  that  they  do  not  recover  more 
than  50  per  cent.  In  the  southern  fields  better  results  are 
claimed,  since  the  cover  is  about  110  ft.  thick  and  all  soft. 

The  slightly  inferior  seams  of  coal  above  or  below  the  Nos. 
5  and  6  seams  are  now  receiving  considerable  thought  as  to 
future  values,  and  for  this  reason  they  are  trying  to  prevent 
roof  movements  by  leaving  sufficient  pillars. 

In  the  Sherrard  field  of  Illinois,  the  recovery  is  reported  at 
90  per  cent.  Here  the  top  and  bottom  are  good  and  conditions 
propitious  for  drawing  pillars.  The  seam  of  coal  is  only  3  ft. 
8  in.  thick,  with  many  clay  veins  running  through  it,  which 
evidently  must  effect  recovery  to  some  extent. 

An  inquiry  sent  into  the  southwest  section  of  Pennsylvania 
shows  a  recovery  of  72.5  per  cent.  Here  10  in.  of  roof  coal  is 
allowed  to  remain  on  account  of  drawslate  and  the  operations 
for  the  past  three  years  have  been  under  the  plant  and  town. 

The  bottom  in  this  mine  is  fireclay  of  rather  soft  character. 
The  rooms  are  driven  from  both  sides  of  entries  on  60-ft.  centers 
and  widened  to  21  ft.,  leaving  39-ft.  pillars  to  be  drawn  by 
machine  and  pick  work.  By  this  method,  the  ultimate  recov- 
ery is  expected  to  show  a  material  increase  over  that  given. 

The  company  reporting  from  Westmoreland  County, 
Pennsylvania,  where  conditions  seem  favorable,  both  in  the 
steam-  and  gas-coal  fields,  shows  a  recovery  of  from  82  to  86 
per  cent  of  the  entire  seam ;  and  expects  the  ultimate  recovery 
to  fall  below  these  figures. 


168 


COAL  MINING  COSTS 


In  Somerset  County,  Pennsylvania,  where  coals  of  the 
Allegheny  series  are  worked,  the  recovery  is  given  as  94.75 
per  cent  of  the  entire  seam.  Here,  excellent  roof  and  bottom 
conditions  prevail,  and  most  room  headings  are  driven  to  the 
limit  before  any  rooms  are  driven  at  all.  Then,  the  rooms 
are  started  at  the  rear  and  pillars  drawn  as  soon  as  these  are 
finished. 

Practically  the  same  conditions  exist  in  the  George's  Creek 
field  in  the  Sewickley  seam,  only  the  recovery  is  reported  as 
97  per  cent.  In  this  same  field,  the  results  obtained  in  the 

TABLE  OF  PRINCIPAL  FACTORS  GOVERNINO 


Per  Cent 

Ultimate 

Period 

5 

Operating  District 
and  State 

of 
Recovery 
of 

Recovery 
Compared 
to 

of 
Opera- 
tion, 

Average 
Height  of 
Seam 

Roof  Coal 
Carried 

Entire 
Seam 

Present 

Years 

i 

Southern  Colorado  

80-90 

Same 

5  to  30 

8'  6" 

0  to  24" 

2 

Colorado  (other  Districts)  .  . 

60-65 

Same 

5  to  30 

8'  6" 

18"  to  24" 

3 

Colorado  (other  Districts)  .  . 

75-80 

Better 

10  to  35 

3'  to  V 

Few  places 

4 

Saginaw  District,  Michigan. 

65,  80-90 

Better 

15 

3' 

1  of  10 

5 

Central  Illinois  

50 

Same 

20 

8' 

None 

6 

Southern  Illinois  

65-70 

Same 

20 

8' 

Yes 

7 

Springfield  District,  Illinois. 

55-75 

Increase  4 

20  to  25 

6'  to  7' 

Some  places 

8 

Franklin,    WiUamson    and 

Saline  Counties,  111  

55-75 

Increase4 

18 

5*' 

Some  places 

.9 

Sherrard  Field,  111  

90 

Same 

20 

3'    8" 

None 

10 

Extreme  Southwest  Section, 

Pennsylvania  

72* 

Better 

3 

T    6" 

10" 

11 

Pennsylvania-Westmoreland 

Co  

84 

Below 

25  to  35 

6'    8" 

None 

11 

Pennsylvania  -Somerset 

Field  

95 

Same 

8 

3'  11" 

None 

13 

Maryland-Georges      Creek 

Field  .  ... 

97 

Same 

12 

3'    0" 

None 

14 

Maryland-Georges      Creek 

Field  

88 

Same 

94 

9'    0" 

18" 

15 

Ohio,  Belmont  County.  .  .  . 

60 

Same 

5'    6" 

None 

16 

Eastern      Ohio,      Harrison 

County  

70-75 

Same 



3'  8"  to  5'  0" 

None 

17 

West  Virginia  

90 

Same 

50 

8' 

Some  places 

18 

Alabama  

19 

Tennessee  

20 

Kentucky  

No  reply 

21 

Kansas  

22 

Iowa                      

1  Length  of  room — 150  feet. 

2  Advocates  retreat  mining. 

3  No.  6  seam. 


MINING  COSTS 


169 


"Big  Vein"  or  Pittsburgh  seam,  show  88  per  cent.  Consider- 
able propping  is  necessary,  owing  to  the  drawslate  and  the 
wild  coal  just  above  it.  The  systems  of  mining  the  coal  in 
this  field  have  changed  from  time  to  time  until  now  headings 
are  driven  9  ft.  wide,  rooms  only  13  ft.  wide  and  the  distance 
between  room  centers  maintained  at  100  ft.,  thus  providing 
against  squeezes.  Under  this  process,  90  per  cent  extraction  is 
expected. 

In  Ohio,  as  well  as  some  parts  of  West  Virginia,  no  attempt 
is  made  to  draw  pillars  at  all.     Rooms  are  driven  25  ft.  wide 

RECOVERY  OF  COAL  IN  DIFFERENT  DISTRICTS 


Nature  of  Top 

Nature  of  Bottom 

System  of 
Mining 

Are  Pillars 
Drawn 

Clay 

Veins 
En- 
countered 

Slate 
Very  soft 

Soft  slate 
Soft  slate 

Room  and  pillar 
Room  and  pillar 

Yes 

Yes 

Dikes 
None 

Sandstone,  poor  shale 

Same  s  s  top 

Room  and  pillar 

Yes 

None 

Black  slate       . 

Fire  clay 

Room  and  pillar 

Where  allowed 

None  i 

Slate,  clod  and  limestone 

Fire  clay 

Panel  system 

None 

None  2 

Sandy  shale 

Fire  clay 

Panel  system 

To  some  extent 

None  3 

Hard  shale 

Fire  clay 

Room  and  pillar 

Where  allowed 

None 

Hard  shale 

Fire  clay 

Room  and  pillar 

Where  allowed 

None 

Blue  rock  and  cap  rock 

Slate  and  sand  rock 

Room  and  pillar 

Yes 

Yes 

18"  to  3'  0  '  draw  slate 

Soft  fire  clay 

Room  and  pillar 

Yes 

None 

Slate 

Fire  clay 

Room  and  pillar 

Yes 

None 

Hard  black  slate 

Limestone 

Room  and  pillar 

Yes 

Very  few« 

Sand  rock 

Sand  rock 

Room  and  pillar 

Yes 

Yes  « 

{Gray  shale,  coal,  } 
dark  shale        J 

Hard  gray  shale 

Room  and  pillar 

Yes 

None 

Slate  and  shale 

Fire  clay 

Room  and  pillar 

None 

None 

10"  firm  slate 

Fire  clay 

Room  and  pillar 

None 

None 

Varies 

Fire  clay 

Room  and  pillar 

Yes 

Yes 

4  Adjustment  of  labor  situation. 
»  Projected  work  adhered  to. 
1  Sewickley  seam. 


170  COAL  MINING  COSTS 

with  8  to  12-ft.  pillars  between.  In  one  of  the  largest  mines 
of  Belmont  County,  rooms  were  driven  from  both  sides  of  the 
headings  and  it  was  no  infrequent  occurrence  to  have  a  ter- 
ritory squeeze  shut,  leaving  considerable  blocks  of  coal  between 
the  ends  of  unfinished  rooms.  In  this  mine,  50  per  cent  would 
approximate  the  recovery. 

In  Harrison  County,  Ohio,  it  is  nearly  as  bad.  The  recov- 
ery is  reported  as  70  to  75  per  cent,  but  the  same  conditions 
exist  in  this  section  as  in  Belmont  County,  excepting  perhaps 
the  driving  of  rooms  both  ways  from  the  same  entry.  The 
Ohio  Mining  Commission  found  in  its  investigations  that  30, 
40  and  as  high  as  50  per  cent  of  coal  is  being  left  underground 
as  pillars  in  that  state. 

There  are  mines  in  West  Virginia  which  show  recovery 
from  85.6  per  cent  to  99.8  per  cent,  the  highest  percentage 
resulting  from  the  fact  that  all  the  work  was  in  the  solid.  The 
average  result  of  the  figures  presented  for  10  mines  showed 
about  92.6  per  cent. 

The  foregoing  figures  reveal  what  is  possible,  at  the  same 
time  showing  what  is  actually,  presumably,  being  accomplished. 

It  would  not  be  proper  to  accept  an  average  of  the  per- 
centages here  given  as  a  fair  maximum,  nor  even  an  average 
of  the  same  field,  as  it  is  unfair  to  compare  ultimate  recovery 
of  mines  now  drawing  to  a  close  with  those  at  the  best  of 
their  production.  No  doubt,  the  systems  under  which  they 
were  inaugurated  were  considered  modern,  but  they  would  not 
be  considered  so  now. 

From  reports  sent  in,  it  is  apparent  that  there  are  five 
factors  limiting  the  possible  recovery  in  these  fields  as  follows : 

1.  Mining  rights  and  public  feeling. 

2.  Roof  and  bottom  conditions. 

3.  Weight  and  character  of  overburden. 

4.  Labor  conditions. 

5.  Market  value  of  the  coal. 

1.  Where  the  mining  rights  do  not  allow  breaking  of  sur- 
face,  the  recovery  naturally  varies  inversely  in  some   ratio 
to  the  overlying  weight. 

2.  Where  roof  and  bottom  conditions  make  it  necessary  to 
recover  as  quickly  as  possible,  market  conditions  will  affect 
recovery  for  pillar  work  of  this  kind  will  not  wait. 


MINING  COSTS  171 

3.  Weight  and  character  of  overburden  require  systematic 
mining  and  competent  supervision. 

4.  It  is  a  matter  of  what  is  next  best  when  unions  insist 
on  conditions  which  increase  both  cost  of  operating  and  loss 
of  coal. 

5.  The  market  value  of  the  coal  dictates  how  far  it  is  pos- 
sible to  go  toward  its  recovery. 

These  points  are  mentioned  because  there  is  a  tendency  to 
compare  straight  figures  of  recovery  without  taking  into  con- 
sideration the  conditions  under  which  they  are  derived.  The 
Ohio  Mining  Commission,  for  instance,  uses  the  mines  and 
operations  at  Gary  for  an  example  of  what  Ohio  should  follow. 
Conditions,  however,  are  so  different  in  these  two  localities, 
that  to  secure  the  same  results  in  recovery  would  require 
several  radical  changes.  Generally  the  roof  in  Ohio  is  poor, 
union  scales  require  rooms  entirely  too  wide  for  economic  pillar 
drawing,  the  general  labor  situation  is  always  more  or  less 
unsettled  and  the  selling  qualities  of  the  coal  are  inferior  to 
those  of  the  Pocahontas  seam  at  Gary. 

Many  of  the  difficulties  incident  to  the  adoption  of  adequate 
conservation  measures  were  described  in  a  paper  presented 
before  the  West  Virginia  Mining  Institute  in  1908.  There  are 
four  factors,  any  one  of  which  is  sufficient  to  cause  a  serious 
loss  in  the  percentage  of  coal  recovered  and  an  increase  in  cost 
of  production,  reducing  the  ultimate  earnings  of  the  property : 

First — Insufficient  or  incompetent  engineering :  Until  very 
recently  the  engineer  was  regarded  by  many  coal  operators  as 
a  luxury  and  an  unnecessary  refinement.  This  unreasonable 
conservatism,  or  prejudice,  still  exists  to  some  extent;  but  the 
rapid  depletion  of  properties  which  have  been  regarded  for 
many  years  as  practically  inexhaustible  is  finally  bringing  the 
operator  to  a  realization  of  the  necessity  for  the  careful  plan- 
ning and  scientific  projection  of  his  mine  by  a  competent  and 
sufficient  engineering  force. 

Second — Incompetent  management:  There  are  some  mine 
managers  or  superintendents  who  produce  better  results  from 
a  poorly  designed  mine  than  others  can  obtain  from  a  first-class 
plant.  Many  managers  have  been  entirely  satisfied  with  a  low 
cost-sheet,  neglecting  the  conditions  both  inside  and  outside 
the  mine,  which  ultimately  resulted  in  an  abnormal  increase  in 


172  COAL  MINING  COSTS 

cost  of  production  or  abandonment  of  valuable  acreages  of 
coal,  in  order  to  maintain  lower  cost.  This  method  would  be 
repeated  until,  finally,  after  millions  of  tons  of  coal  had  been 
ruthlessly  buried,  which  could  have  been  recovered  by  a 
prudent  and  careful  manager  without  materially  affecting  his 
cost  of  production,  he  was  brought  to  a  sudden  realization  of 
an  unnecessarily  high  cost  and  a  diminished  tonnage,  result- 
ing in  loss  of  prestige  and  position  for  him  and  thousands  of 
dollars  to  the  owners.  The  necessity  for  a  mine  manager  to 
be  familiar  with  not  only  the  outside,  but  also  the  inside,  con- 
dition of  the  property,  either  personally  or  through  tried  and 
competent  assistants,  cannot  be  underestimated.  Unless  he  is 
thoroughly  familiar  with  these  conditions,  how  can  he  know 
if  the  individual  mine  superintendent  or  foreman  is  giving 
him  the  desired  results  of  low  cost  with  maximum  recovery, 
and  the  best  conditions  for  a  continuation  of  that  low  cost 
and  recovery? 

Third — Unfavorable  labor  conditions :  The  employment  of 
unskilled  miners  renders  it  impossible  to  obtain  a  good  recov- 
ery. Strikes  and  shut-downs  have  often  interfered  with  the 
application  of  economic  methods  of  extracting  coal.  Labor 
unions  sometimes  insist  on  conditions  which,  while  operating 
for  the  convenience  of  the  miner,  increase  the  expense  or  induce 
an  unnecessary  loss  of  coal  to  the  operator.  For  instance,  in 
some  districts  of  the  country  the  track  must  be  laid  in  the 
center  of  the  room,  with  the  gob  on  either  side.  Few  or  none 
of  the  pillars  are  recovered,  and  many  acres  of  coal  have  been 
lost  through  squeezes  and  creeps  because  of  wide  rooms  and 
small  pillars.  Nor  is  interference  with  the  methods  of  mining 
the  only  manner  in  which  the  unions  sometimes  operate  against 
the  maximum  recovery  of  coal.  There  have  been  men  dis- 
charged for  incompetency  who  were  reinstated  by  the  manage- 
ment on  demand  of  the  union.  It  is  thus  that  discipline,  which 
is  such  an  important  factor  in  the  economic  administration  of 
a  mine  with  a  view  to  the  best  ultimate  results,  is  destroyed. 

Fourth — Impatience  of  owners  for  quick  returns  on  invest- 
ment: The  deleterious  effect  on  the  economic  development  of 
a  property  by  the  demand  of  the  investor  for  an  immediate 
profit  can  hardly  be  overestimated.  Many  a  rich  and  valuable 
property  has  been  irreparably  damaged  by  the  insistence  of 


MINING  COSTS  173 

owners  for  immediate  and  large  profits  before  its  proposed 
economic  development  has  been  fairly  launched.  This  impa- 
tience and  greed  has  at  times  resulted  in  the  changing  of  slow, 
but  good  plans  of  development  for  bad  ones,  and  the  poor  results 
thus  obtained  were  further  accentuated  in  later  years  by  the 
demand  for  big  tonnage,  thereby  causing  the  loss  of  millions 
of  tons  of  coal  and  thousands  of  dollars  to  the  investor.  In- 
deed, this  demand  for  large  tonnage  from  a  poorly  developed 
mine  has  been  the  greatest  factor  in  encouraging  careless 
methods  of  recovering  coal  both  from  rooms  and  pillars,  and 
it  is  no  exaggeration  to  state,  that  the  abandoning  of  many 
acres  of  good  coal,  which  could  have  been  recovered  by  more 
thorough  and  known  methods,  can  be  traced  directly  to  this 
cause. 

Use  of  the  longwall  system  to  effect  conservation.— The 
adoption  of  the  longwall  system  of  mining,  where  possible,  will 
be  the  ultimate  solution  to  obtaining  the  maximum  recovery  of 
coal  and  the  subject  is,  therefore,  one  for  the  serious  considera- 
tion of  the  coal  economist. 

To  obtain  a  comparison  of  the  results  of  the  different  sys- 
tems a  1000-acre  tract  of  land  with  6  ft.  of  coal  lying  at  a 
depth  of  400  ft.  will  be  assumed.  This  depth  is  taken  to  make 
allowance  for  the  possibilities  of  the  longwall  system  on  sur- 
face caving,  one  of  its  chief  disadvantages. 

A  conservative  estimate  of  the  recovery  to  be  obtained 
from  such  a  tract  under  present  systems  of  mining  would  be 
about  60  per  cent.  The  total  tonnage  in  this  acreage  would  be 
9,680,000  short  tons,  60  per  cent  of  which  would  be  5,800,000 
tons  which  means  a  loss  of  nearly  4,000,000  tons.  At  a  value  of 
$1  per  ton  this  would  mean  a  loss  in  the  national  wealth  of 
$4,000,000. 

In  removing  the  entire  seam  some  damage  has  perhaps 
resulted  to  the  farmer,  but  not  much,  certainly,  at  a  depth  of 
400  ft.,  and  nothing  beyond  easy  repairs.  Where  a  total  extrac- 
tion has  been  effected  on  a  5-ft.  seam  having  100  ft.  cover  in 
certain  of  the  Pennsylvania  fields,  the  break  has  extended  to 
the  grass  roots.  Where  this  extraction  has  "been  several  acres 
in  extent  and  the  break  has  been  general  over  the  entire  area 
at  once,  it  cannot  be  said  that  any  appreciable  damage  has 
resulted.  In  this  instance  there  was  no  packing  whatever,  as 


174  COAL  MINING  COSTS 

would  ordinarily  be  the  case  in  longwall  workings.  The  con- 
ditions were  simply  100  ft.  of  easily  breaking  roof  with  a  clean 
fall  of  5  ft. 

Taking  our  previous  example  again,  of  6  ft.  of  coal  at  a 
depth  of  400  ft.  worked  by  the  longwall  system  and  packed 
carefully,  the  result  would  by  no  means  be  so  serious.  Further- 
more, it  should  not  be  forgotten  that  there  are  approximately 
1000  sq.  mi.  of  coal  in  which  the  seams  are  600  ft.  or  more 
below  the  surface.*  At  this  latter  depth  it  is  doubtful  if  the 
strata  would  break  to  the  surface,  for  it  has  been  shown  in  the 
British  mines  that  at  a  depth  of  700  or  800  ft.  work  can  be 
carried  on  successfully  under  the  sea. 

This  proves  that  in  average  strata  the  highest  fall  reaches 
an  apex  well  below  that  distance,  and  the  miner  who  has  had 
extensive  experience  in  pillar  drawing  and  is  familiar  with 
the  quick  oblique  line  that  the  ragged  rock  edges  traverse 
toward  a  common  juncture,  will  place  the  safety  point  much 
lower. 

Turning  again  to  our  example  of  a  1000-acre  tract,  and 
assuming  this  to  have  been  worked  by  the  longwall  system 
which  has  resulted  in  certain  damage  to  the  surface,  a  com- 
parison of  this  surface  damage,  with  the  additional  extraction 
obtained,  may  be  made.  Taking  the  mine  on  a  royalty  basis 
of  5c.  a  ton,  the  farmer  has  received  $193,600  above  what  he 
would  have  received  by  the  60  per  cent  pillar-and-room  method 
of  working.  This  amounts  to  $193  per  acre,  or  about  twice 
as  much  as  the  average  farm  value  of  land.  It  should  also  be 
remembered  that  the  expense  of  tipple  erection,  compressors 
and  power  plants  and  the  general  surface  arrangement  of  a 
mine  opened  to  develop  a  tract  of  1000  acres,  is  the  same  for 
60  per  cent  extraction  as  for  100. 

It  is  said  that  in  longwall  working  continuous  operation  is 
necessary,  in  order  to  properly  control  the  roof,  but  this  applies 
nearly,  if  not  quite  as  well,  to  room-and-pillar  work,  particu- 
larly when  the  mine  contains  water.  The  statement  of  a  cer- 
tain large  Western  operator,  a  man  who  is  both  practical  and 
theoretical,  may  be  taken,  apropos  of  this.  He  changed  a  num- 
ber of  his  mines  to  the  longwall  system,  and  during  a  strike 

*See  Twenty-second  Annual  Report,  U.  S.  Geological  Survey,  page  178. 


MINING  COSTS  175 

lasting  over  a  period  of  about  five  months  and  a  half,  all  of 
his  mines  were  shut  down.  On  the  resumption  of  operations 
he  had  careful  records  kept  of  the  relative  cost  of  opening  the 
longwall  and  the  room-and-pillar  mines.  These  records  showed 
the  cost  of  opening  the  room-and-pillar  mines  to  be  nine  times 
that  of  the  longwall. 


SECTION    II 
SHAFT  SINKING 

The  investment  involved  in  shaft  sinking  is  heavier  than 
that  encountered  with  any  other  improvement.  Mistakes  can 
be  neither  rectified  nor  lived  down.  Time  and  first  cost  being 
the  essence  of  the  opening  of  a  new  property,  important  fea- 
tures are  often  sacrificed  underground  instead  of  on  the  sur- 
face, where  future  remodeling  is  practicable.  It  is,  therefore, 
obvious  that  the  preliminary  engineering  and  estimating  rela- 
tive to  a  shaft  operation  deserve  serious  study. 

In  planning  a  shaft  mine  opening  the  following  points  come 
up  for  consideration:  Avoidance  of  all  unnecessary  narrow- 
work  at  the  bottom  increasing  the  risk  of  loss  from  squeeze, 
compliance  with  all  present  and  possible  future  legislation, 
minimum  first  and  maintenance  cost,  the  connecting  up  of  the 
two  shafts  in  the  shortest  time  in  order  that  a  regular  circula- 
tion of  air  may  be  obtained  and  the  restrictions  of  the  law 
limiting  the  number  of  men  allowed  in  the  mine  before  this  is 
done  complied  with.  The  arrangement  of  the  work  should  be 
such  that  during  development  cars  may  be  placed  with  the 
utmost  facility,  that  mining  machines  may  be  employed  and 
steel  timbering  placed  as  the  entries  advance;  that  when 
operations  begin,  cars,  supplies,  waste,  men,  water,  air,  dam- 
aged equipment,  etc.,  may  be  handled  with  safety,  economy 
and  speed;  that  protection  is  secured  against  coal-dust  explo- 
sions, mine  fires,  flooding,  freezing,  etc. 

Operations. — The  universal  method  of  shaft  sinking  in  rock 
is  to  drill  a  number  of  holes  in  the  bottom,  charge  them  with 
dynamite  and  shoot  them,  and  to  load  the  broken  rock  by 
hand  into  buckets  which  are  then  hoisted  out.  When  all  the 
loose  rock  has  been  removed  the  process  is  repeated. 

Shafts  are  drilled  on  the  "  center-cut "  principle.  Eight  or 
ten  holes  are  drilled  on  a  slant,  separated  at  the  top  but  con- 
verging, thus  forming  a  wedge  known  as  the  "sump."  "Re- 

176 


SHAFT  SINKING 


177 


liever,"  or  bench,  holes  are  drilled  back  of  the  sump  holes, 
each  row  being  more  nearly  vertical;  the  end  or  outside  holes 
point  slightly  away  from  the  vertical  and  toward  the  wall 
line  of  the  shaft.  The  sump  is  first  shot  and  the  broken  rock 
removed  or  "  mucked "  out,  forming  a  cavity  into  which  the 
bench  rounds  can  be  successively  shot.  All  muck  should  be 
removed  before  each  succeeding  round  is  shot. 

Two  systems  of  drilling  and  mucking  exist.  In  the  first 
the  holes  for  the  entire  cut — sump  and  benches — are  drilled 
at  one  time,  the  sump  is  shot,  and  then  the  benches  as  required. 
In  the  second,  the  sump  only  is  drilled  and  shot,  and  the 
benches  are  drilled  while  the  sump  is  being  mucked.  The  first 
plan  is  particularly  applicable  to  small  shafts  and  to  circular 
shafts ;  a  rectangular  or  elliptical  shape  is  needed  to  give  room 
for  simultaneous  drilling  and  mucking. 

Fumeless,  or  gelatine,  dynamite  should  in  all  cases  be  used 
for  underground  work.  The  fumes  from  ordinary  glycerine 
dynamite  make  it  impossible  for  the  men  to  get  back  to  work 
promptly  after  a  shot.  The  strength  of  the  dynamite  used 
depends  on  the  character  of  the  rock,  but  40-per-cent  and  60- 
per  cent  gelatine  are  the  most  common  strengths  used. 

The  number  and  depth  of  the  holes  and  the  quantities  of 
powder  loaded  vary  so  greatly  with  the  size  of  the  shaft  and 
the  nature  of  the  rock  that  no  general  rules  can  be  stated.  The 
systems  actually  used  at  several  shafts  were  as  follows: 

Shaft  13  X  26  ft.,  through  Western  Pennsylvania  coal 
measures:  Shale,  slate,  and  limestone;  horizontal  stratifica- 
tion; 40-per-cent  gelatine: 


Number 

Depth, 
Feet 

Inclina- 
tion with 
Vertical, 
Degrees 

Loaded 
with 
Pounds 

Sump   . 

8 

10 

45 

4 

Relievers  
Benches  

8 

8 

8 
8 

30 
0 

3 

2i 

End 

8 

8 

10  back 

2* 

Total  charge  

96 

Average  gain  per  cut,  6  ft. 

Average  gain  per  week  of  19  shifts,  24  ft.  (no  timber). 

Mucking  and  drilling  simultaneous;  2  drills  used  on  1  bar,  double. 


178 


COAL  MINING  COSTS 


Shaft  14  X  48  ft.,  through  anthracite  measures :    Red  sand- 
stone; stratification  horizontal;  40-per-cent  gelatine: 


Number 

Depth, 
Feet 

Inclina- 
tion, 
Degrees 

Loaded 
with, 
Pounds 

Sump               

8 

10 

45 

5 

Relievers 

8 

8 

30 

4 

Benches       

24 

8 

10 

3 

End 

8 

8 

10  back 

3 

Total  charge  per  round  

168 

Average  gain  per  cut,  6  ft. 

Average  gain  per  week  of  18  shifts,  16  ft. 

Mucking  and  drilling  simultaneous;  2  drills  used  on  1  bar. 

Shaft  10  X  22  ft.,  through  quartz  conglomerate  (Shawan- 
gunk  grit) ;  horizontal  stratification,  but  very  few  bedding 
planes;  60-per-cent  gelatine: 


Number 

Depth, 
Feet 

Inclina- 
tion, 
Degrees 

Loaded 
with, 
Pounds 

Sump   . 

8 

10 

45 

31 

Sump 

4 

8 

o 

31 

Relievers  

8 

9 

30 

2£ 

Benches               

8 

8 

o 

2 

End 

8 

8 

10  back 

2 

Total  charge  per  round  

94 

Average  gain  per  cut,  5£  ft. 

Average  gain  per  week  of  20  shifts,  22  ft. 

Mucking  and  drilling  simultaneous;  5  drills  used  on  2  bars. 

The  four  additional  sump  holes  shown  were  used  on  account 
of  extra  hardness  of  the  rock. 

Shaft  elliptical,  19  ft.  4  in.  X  33  ft.,  through  West  Virginia 
coal  measures:  Hard  gray  sandstones;  40-per-cent  gelatine; 
horizontal  stratification : 


SHAFT  SINKING 


179 


Number 

Depth, 
Feet 

Inclina- 
tion, 
Degrees 

Loaded 
with 
Pounds 

Sump   

10 

12 

45 

5 

Relievers     

8 

10 

30 

4 

Benches 

14 

10 

10 

4 

End      

6 

10 

10  back 

3 

Total  charge  per  round 

156 

Average  gain  per  cut,  8  ft. 

Average  gain  per  week  of  20  shifts,  18  ft. 

Mucking  and  drilling  simultaneous;  3  drills  used  on  1  long  bar,  1  short  bar. 

Shaft  circular,  17  ft.  diameter,  through  Hamilton  and  Mar- 
cellus  shales:  Rock  distorted;  stratification  irregular;  but 
about  45  deg. ;  60-per-cent  gelatine : 


Number 

Depth, 
Feet 

Inclina- 
tion, 
Degrees 

Loaded 
with 
Pounds 

Sump     

6 

8 

45 

2i 

Relievers 

8 

6 

20 

14 

Rib   

16 

6 

10  back 

1 

Total  charge  per  round 

43 

Average  gain  per  cut,  5|  ft. 

Average  gain  per  week  of  19  shifts,  33  ft. 

All  drilling  on  one  shift,  mucking  on  two  shifts;  5  drills  used  on  5  tripods. 

While  the  hand  drilling  has  been  displaced  almost  entirely 
by  power  driven  drills,  there  are  still  occasions  when  a  small 
job  does  not  justify  the  installation  of  power  equipment  and 
it  is  more  economical  to  resort  to  hand  drilling.  This  is  par- 
ticularly so  in  foreign  countries  where  labor  costs  are  fre- 
quently quite  low. 

The  customary  procedure  is  the  use  of  a  1-in.  drill,  turned 
by  one  man  and  struck  by  one  or  two  others  with  8-lb.  hammers. 
Two  strikers  should  always  be  used  where  practicable  as  they 
can  obviously  drill  twice  as  fast  as  a  single  striker  at  three- 
fourths  the  cost,  Three  capable  men  can  drill  1^-in.  holes 
in  hard  sandstone  at  the  rate  of  2  ft,  per  hour. 


180  COAL  MINING  COSTS 

In  soft  material,  churn  drills  6  to  12  ft.  long  with  a  bit  at 
each  end  are  sometimes  used  with  satisfactory  results.  These 
drills  are  sometimes  weighted  to  give  additional  striking  power 
and  they  are  usually  operated  by  two  or  three  men. 

Shaft  sinking  is  usually  carried  on  24  hr.  a  day.  The  inside 
work  is  done  by  three  shifts  of  men  working  8  hr.  each,  the 
outside  by  three  8-hr,  or  two  12-hr  shifts.  The  12-hr,  outside 
shift  is  customary  in  the  coal  fields;  elsewhere,  the  8-hr,  shift 
for  every  one  is  prevalent.  Shifts  are  usually  changed  at  7  a.m. 
and  3  and  11  p.m.,  sometimes  an  hour  later.  The  men  are  given 
20  min.  for  lunch  in  the  middle  of  each  shift. 

Wages  vary  with  the  locality,  but  in  general  men  are  paid 
better  for  drilling  and  mucking  in  a  shaft  than  in  any  other 
kind  of  rock  excavation.  On  account  of  the  high  wages  paid 
in  America  machine  drilling  is  universal,  and  the  shifts  are 
limited  to  the  number  of  men  that  can  be  worked  to  the  best 
advantage.  Speed  is  not  attempted  at  the  expense  of  efficiency. 
In  South  Africa,  on  the  other  hand,  Kaffir  labor  is  cheap,  hand 
drilling  is  usual,  and  as  many  men  are  worked  as  the  shafts 
will  hold. 

:The  great  clepth  of  the  shafts  on  the  Rand  makes  the  high- 
est possible  speed  desirable,  even  at  an  increased  cost.  In 
both  countries  speed  is  increased  without  an  increase  of  cost 
by  the  payment  of  a  bonus  to  the  sinkers  as  a  reward  for 
additional  progress. 

The  size  of  the  shifts  for  any  given  shaft  depends  upon  the 
number  of  drills  required  and  upon  the  experience  and  ability 
of  the  sinkers  obtainable.  With  first-class  men,  the  men  on 
each  shift  for  a  13  X  26  ft.  shaft  would  be  as  follows,  wages 
as  of  1909 : 

Inside  men,  8  hr. :  One  shiftboss,  at  $3;  two  drillers,  at 
$2.75;  two  helpers,  at  $2.50;  six  muckers,  at  $2.25. 

Outside  men,  12  hr. :  One  engineer;  one  head  tender;  three 
car-men  on  dump ;  one  firemen ;  one  compressor  man. 

General  outside,  10  hr.:  One  foreman;  one  mechanic;  two 
carpenters  (on  timber)  ;  one  blacksmith  and  helper. 

A  17-ft.  circular  shaft  would  require: 

Drilling  shift:  One  shiftboss;  five  drillers;  five  helpers; 
one  extra  man. 

Mucking  shift;    One  shiftboss;  nine  muckers. 


SHAFT  SINKING  181 

Outside:    Same  as  above. 

In  South  African  shafts,  which  are  usually  9  X  26  ft.,  drill- 
ing is  always  done  by  hand  and  each  shift  consists  one  white 
shiftboss  and  35  Kaffir  laborers  who  drill  or  muck  as  may  be 
required. 

Thorough  organization  is  essential  to  progress  and  economy. 
Each  man  must  know  his  place  and  take  it  without  losing  time 
in  getting  started.  Any  system  that  prevents  systematic  work 
is  fatal  to  economy. 

Circular  or  Rectangular. — From  a  construction  standpoint 
the  circular  or  elliptical  and  rectangular  types  are  equally 
feasible,  and  the  choice  depends  upon  the  cost.  In  several 
cases  a  compromise  has  been  effected  by  shaping  the  shaft  as 
a  quadrilateral  with  sides  formed  of  circular  arcs. 

For  a  single  compartment  air-shaft  the  circular  shape  is  in 
every  way  the  most  desirable,  not  only  because  the  circular 
shaft  is  cheaper  to  sink  than  a  square  shaft  of  equal  area,  but 
also  because  a  circular  ring  of  plain  concrete  is  the  strongest 
lining  possible  with  a  given  amount  of  material. 

In  the  case  of  a  shaft  with  two  or  more  compartments,  the 
selection  of  the  most  economical  shape  requires  some  calcula- 
tion. At  first  sight  it  would  seem  that  a  simple  rectangular 
shaft  surrounded  by  a  concrete  wall  only  thick  enough  to  be  as 
strong  as  the  usual  timber  lining,  would  be  a  satisfactory,  as 
well  as  a  cheap,  shape,  but  this  is  not  the  case.  A  concrete 
lining,  even  when  provided  with  weep  holes,  must  resist  some 
hydrostatic  pressure ;  a  timber  lining  has  none  to  resist.  Fur- 
thermore, permanent  weep  holes  are  most  undesirable;  the 
concrete  should  exclude  the  water  entirely,  and  hence  must  be 
designed  to  bear  very  great  pressure  at  considerable  depth. 
Just  what  amount  the  theoretical  pressure  is  reduced,  by  the 
adhesion  of  the  concrete  to  the  shaft  walls  and  by  the  block- 
ing of  the  fissures  with  grout,  cannot  be  calculated.  In  solid 
rock,  where  the  water  enters  in  well-defined  springs,  the  proper 
grouting  of  the  springs  will  relieve  the  lining  of  all  pressure. 
In  very  seamy  rock,  on  the  other  hand,  the  lining  may  have  to 
bear  practically  the  full  hydrostatic  pressure. 

In  order  to  compare  the  costs  of  the  different  shapes,  let  us 
consider  in  detail  three  designs  for  a  shaft  with  two  7  X  10  ft. 
hoistways  and  an  airway  with  an  area  of  100  sq.  ft.  As  the 


182 


COAL  MINING  COSTS 


QUANTITIES  AND  COSTS  OF  RECTANGULAR  SHAFT 


Depth  in  feet 

20 

50 

100 

150 

2ftft 

Total  thickness  of  lining  in  inches  .  .  . 

14 

21 

28 

34 

MHI 

39 

Quantities  per  linear  foot: 

Concrete  to  neat  line  in  cubic  yards. 

3.90 

5.70 

7.60 

9.30 

10.70 

Concrete  actual  in  cubic  yards  

5.80 

7.70 

9.70 

11.50 

13.00 

Excavation  to  neat  line  in  cubic 

yards  

12.80 

14.60 

16.50 

18.20 

19.70 

Excavation  actual  in  cubic  yards  .  . 

14.70 

16.60 

18.60 

20.40 

22.00 

Weight  reinforcing  steel  in  pounds. 

256 

443 

650 

845 

1030 

Cost  per  linear  foot: 

Forms                            

$25.00 

$25  .  00 

$25  .  00 

$25.00 

•eof)  nn 

Concrete  at  $5  cubic  yard    

29.00 

38.50 

48.50 

57.50 

O~«>  .  UU 
«K     Oft 

Excavation  (see  note  *)  

49.60 

53.20 

57.00 

60.40 

oo  .  uu 

AQ     Aft 

Reinforcing  steel  at  $0.02  pound.  . 

5.10 

8.90 

13.00 

16.90 

OO  .  ^U 

20.60 

Total                        

$108.70 

$125.60 

$143.50 

$159.80 

$174   OO 

«IP  1  1  1  .  UU 

QUANTITIES  AND  COST  OF  ELLIPTICAL  SHAFT 


Depth  in  feet,  0  to  

100 

150 

200 

250 

300 

400 

Thickness  of  lining  in  inches,  ends  .... 

12 

12 

12 

12 

12 

12 

Thickness  of  lining  in  inches,  sides  .... 

12 

18 

24 

29 

34 

42 

Quantities  per  linear  foot: 

Concrete  to  neat  line,  cubic  yards.  . 

2.60 

3.40 

4.30 

5.00 

5.70 

6.80 

Concrete  actual  in  cubic  yards  

4.40 

5.20 

6.10 

6.80 

7.50 

8.60 

Excavation  to  neat  line  in  cubic  yards. 

$15.20 

$16.00 

$16.90 

$17.60 

$18.30 

$19.40 

Excavation  actual  in  cubic  yards.  .  . 

17.00 

17.80 

18.70 

19.40 

20.10 

21.20 

Costs  per  linear  foot: 

15  00 

15  00      ifi  on 

15  00 

15  00 

15  00 

Concrete  at  $5  cubic  yard  

22.00 

26.00 

30.50 

34  00 

37  00 

43  00 

54  40 

56  00 

57  80 

59  20 

60  60 

62  80 

Total 

$91.40 

$97.00 

$103.30 

$108.20 

$113.10 

$120.80 

QUANTITIES  AND  COSTS  OF  QUADRILATERAL  SHAFT 


Depth  in  feet  0  to 

100 
12 

2.70 
4.50 
14.90 
16.70 

$15.00 
22.50 
53.80 

150 
19 

4.40 
6.30 
16.60 
18.50 

$15.00 
31.50 
57.20 

200 
26 

6.20 
8.20 
18.40 
20.40 

$15.00 
41.00 
60.80 

250 
32 

7.90 
10.00 
20.10 
22.20 

$15.00 
50.00 
64.20 

300 
39 

9.90 
12.10 
22.10 
24.30 

$15.00 
60.50 
68.20 

400 
52 

13.90 
16.20 
26.10 
28.40 

$15.00 
81.00 
76.20 

Thickness  of  lining  in  inches  

Quantities  per  linear  foot: 
Concrete  to  neat  line  in  cubic  yards  . 
Concrete  actual  in  cubic  yards  
Excavation  to  neat  line  in  cubic  yards. 
Excavation  actual  in  cubic  yards  .  .  . 
Costs  per  linear  foot: 
Forms                     

Concrete  at  $5  cubic  yard  

Excavation  (see  note  *)  

Total              

$91.30 

$103.70 

$116.80 

$129.20 

$143.70 

$172.20 

*  Cost  of  excavation  figured  on  basis  of  $4  per  cubic  yard  for  section  containing  12  yards 
per  linear  foot;  additional  excavation  at  $2  per  cubic  yard.  Thus  cost  of  16  cubic  yard 
section  =  12  X  $4 +4  X  $2  =  $56. 


SHAFT  SINKING 


183 


whole  area  of  a  hoist  shaft  is  ordinarily  used  for  the  passage 
of  air;  the  size  of  the  air  compartment  may  be  reduced  if  the 
rest  of  the  shaft  is  enlarged;  the  airway  must  however  be 
large  enough  to  contain  pipes  and  ladders  and  to  provide  in 
addition  an  ample  passage  for  air  if  the  hoistways  are  tem- 
porarily closed. 

Let  us  assume  a  minimum  thickness  of  12  in.  of  concrete 
for  a  water-tight  lining ;  also  that  in  each  case  the  lining  carries 
the  entire  hydrostatic  pressure ;  then  the  specifications  for  the 
three  forms  of  shafts  will  be  as  follows: 

Rectangular  Shaft.— Fig.  1.  Two  hoistways  7  X  10  ft.,  one 
airway  10  X  10  ft.  Ten-inch  concrete  dividing  walls  in  place  of 
buntons.  Extreme  inside  dimensions  10  X  25  ft.  8  in.  Area 


.1  2 

! 
j'-~ 

.  >  .-/o'-0".  . 

so     * 

^,'-0- 

1 

. 


FIG.  1. — Rectangular  concrete-lined  shaft. 

airway  100  sq.  ft.,  total  clear  area  240  sq.  ft.  Thickness  of 
lining  at  any  point  made  equal  to  depth  of  simple  beam  of  10  ft. 
span  required  to  sustain  hydrostatic  pressure  at  that  point. 
Resisting  moment  and  weight  of  reinforcement  calculated  by 
Johnson's  formula,  factor  of  safety  3.  (Ultimate  tensile 
strength  of  steel  65,000  Ib.  per  square  inch,  compressive  strength 
of  concrete  in  beam  2500  per  square  inch.)  Reinforcing  steel 
set  3  in.  inside  of  face  of  wall. 

Cost  of  forms,  given  in  the  accompanying  table,  includes  cost 
of  forms  for  dividing  walls,  and  is  therefore  greater  than  the 
cost  in  the  elliptical  shafts. 

Excess  of  actual  over  theoretical  quantity  of  excavation  is 
estimated  as  15  per  cent  for  28-ft.  shaft.  This  excess  increases 
with  the  length  of  the  shaft  only,  as  the  ends  are  drilled  to  line. 

Elliptical  Shaft. — Fig.  2.  Extreme  inside  dimensions  16  X 
27  ft.  Area  of  airway,  78  sq.  ft.  Total  clear  area,  allowing  for 
10-in.  buntons,  304  sq.  ft. 


184 


COAL  MINING  COSTS 


Strength  of  lining  calculated  on  the  assumption  that  the 
stress  in  the  elliptical  cylinder  at  any  point  is  equal  to  that 
caused  in  a  circular  cylinder  with  a  radius  equal  to  the  radius 
of  curvature  of  the  ellipse  at  the  given  point,  by  the  same 
hydrostatic  pressure  acting  upon  it.  The  lining  is  therefore 
made  thicker  at  the  sides  than  at  the  ends. 

To  prove  this  proposition  assume  the  lining  to  be  constructed 
of  a  number  of  small  portions,  each  the  arc  of  a  circle.  The 
stress  in  each  portion  caused  by  the  hydrostatic  pressure  of  the 
film  of  water  between  it  and  the  rock  is  directly  proportional 
to  the  radius,  and  the  thickness  of  each  section  should  therefore 
be  made  proportional  to  the  radius.  Considering  any  portion, 


FIG.  2. — Elliptical  concrete-lined  shaft. 

as  o-&,  Fig.  4,  the  skewback  toward  the  side  of  the  ellipse  is 
formed  entirely  by  the  adjoining  portion,  while  the  skewback 
toward  the  end  is  formed  partly  by  the  adjoining  portion  and 
partly  by  the  rock.  If  the  number  of  circular  portions  is  in- 
definitely increased,  the  unbalanced  end  thrust  of  each  will  be 
taken  up  by  the  irregularities  of  the  rock. 

Ultimate  compressive  strength  of  concrete,  3000  Ib.  per  square 
inch ;  factor  of  safety,  3. 

Excess  of  actual  over  theoretical  excavation  assumed  as  12 
per  cent  for  smallest  section.  As  the  length  of  the  shaft  does  not 
vary,  this  excess  is  constant. 

Quadrilaterial  Shaft. — Fig.  3.  Inside  dimensions,  16  X  24  ft. 
8  in.  Radius  of  ends  and  sides,  23  ft.  Area  of  airway,  94  sq.ft. 
Total  clear  area,  allowing  for  10-in.  buntons,  294  sq.  ft. 

For  calculating  stresses,  sides  and  ends  are  considered  as 


SHAFT  SINKING 


185 


portions  of  a  46-ft.  circular  cylinder.  Ultimate  compressive 
strength  of  concrete  3000  Ib.  per  square  inch,  factor  of  safety,  3. 
Excess  of  actual  over  theoretical  quantity  of  excavation 
assumed  to  be  12  per  cent  for  minimum  length  and  to  increase 
with  the  length. 


FIG.  3. — Quadrilateral  concrete-lined  shaft. 


Germany  is  committed  to  the  circular  or  elliptical  form  of 
shaft,  the  German  engineers  being  of  the  opinion  that  the 
square  or  rectangular  form  is  more  expensive  due  to  the  extra 
work  involved  in  excavating  and  keeping  the  corners  squared 
up. 


FIG.  4. — Design  of  the  elliptical  form  of  shaft. 

It  is  evident,  that  assuming  the  same  hoisting  capacity  in 
either  form  of  shaft,  the  excess  area,  which  makes  ventilation 
possible,  should  be  the  same  in  either  a  circular  or  a  rectangular 
shaft.  A  circular  shaft  of  20  ft.  net  diameter  would  be  roughly 


186  COAL  MINING  COSTS 

equivalent  to  a  rectangular  shaft  12  X  20  ft.  English  mining 
engineers  claim  that  the  cost  of  lining  is  as  5  to  9  in  favor  of 
circular  shafts,  and  it  is  generally  conceded  that  where  great 
pressure  is  encountered  the  circular  form  is  the  only  safe  one. 
A  circular  shaft,  when  once  properly  lined  with  iron  or 
masonry,  is  a  permanent  affair,  while  timber  lining  under  the 
best  conditions  cannot  be  expected  to  last  more  than  18  or 
20  yr.  and  rarely  more  than  15  yr.  It  is  also  well  known  that 
for  a  given  area,  a  circular  shaft  presents  less  rubbing  sur- 
face, or  resistance  to  the  passage  of  the  ventilating  current, 
and  the  segments  at  the  side  of  the  cages  furnish  space  for 
this  air  current  without  additional  enlargement  of  the  shaft. 

The  principal  arguments  advanced  for  rectangular  shafts 
are  that  less  material  needs  to  be  removed  for  a  given  cage 
space,  and  that  in  sinking,  the  permanent  lining  is  at  once 
put  in  place,  as  the  work  progresses. 

It  is  probable  that  the  matter  of  keeping  a  shaft  in  aline- 
ment  during  construction,  by  either  method,  is  largely  a  matter 
of  experience  and  character  of  labor  employed. 

While  engineers  have  claimed  cheapness  of  construction  as 
an  argument  for  both  forms  of  shaft,  the  data  at  hand  would 
indicate  that  the  circular  shaft  may  be  excavated  fully  as 
cheaply  as  the  rectangular,  but  costs  more  to  line;  while  on 
the  other  hand  the  upkeep  and  repairs  on  a  circular  shaft 
properly  constructed,  are  very  much  less  than  on  a  rectangular 
shaft  of  same  capacity,  and  the  danger,  especially  in  deep 
mines  or  in  quicksand,  is  very  materially  reduced,  and  water 
much  more  easily  kept  out,  effecting  a  saving  in  pumping. 

Equipment. — A  two-stage  compressor  is  the  best  for  shaft 
sinking.  With  a  steam  consumption  of  45  Ib.  per  i.h.p.  at  half 
cut-off  the  simple  compressor  has  for  each  indicated  horse- 
power, a  capacity  of  5  cu.  ft.  of  free  air  per  minute  compressed 
to  100  Ib.,  while  the  two-stage  machine  will  deliver  15  per  cent 
more  air  with  the  same  steam  consumption.  For  500  cu.  ft. 
of  free  air  per  minute,  the  saving  of  the  two-stage,  over  the 
simple  type,  will  amount  to  15  per  cent  X  i  X  500  cu.  ft.  X  45 
=675  Ib.  steam  or  150  Ib.  of  coal  per  hour.  With  the  com- 
pressor operating  to  capacity  20  hr.  a  day,  6  days  in  the  week, 
the  saving  in  three  months  with  coal  at  $4.50  per  ton  would  thus 
be  $525. 


SHAFT  SINKING  187 

The  cost  of  a  plant  for  a  single  shaft,  assuming  a  depth  of 
about  500  ft.  and  a  moderate  inflow  of  water,  say  30  or  40  gal. 
a  min,  was  estimated  in  1909  as  follows : 


Sinking  engine $1,000 

Two  80-hp.  boilers  and  setting 1,800 

Pipe  and  auxiliaries 500 

150-hp.  heater 300 

14-in.  compressor 1,750 

Three  drills  and  steel 1,000 

Shaft  bar  and  clamps    '. 100 

Derrick 400 

Head-frame 500 

Two  buckets 150 

Rope 150 

Buildings 500 

Dump  cars  and  rail 300 

Electric  plant,  10  kilowatts 750 

Two  pumps 500 

Small  tools..  500 


Total $10,200 

These  figures  are  based  on  the  cost  of  new  machinery,  and 
are  large  enough  to  include  the  necessary  accessories.  The 
cost  of  erecting  and  dismantling  such  a  plant  will  be  from 
$1000  to  $2000,  depending  on  location,  labor  conditions,  etc. 

Some  of  the  most  serious  pumping  problems  in  shaft  sink- 
ing in  this  country  have  been  encountered  in  the  Pocahontas 
field  along  the  Norfolk  &  Western  Ry.  Except  in  two  instances, 
sinking  has  been  very  expensive  in  this  territory  owing,  it  is 
thought  to  the  fact  that  the  water-bearing  rock  is  a  very  hard 
sandstone  carrying  a  considerable  volume  of  water.  When  a 
shaft  is  started  beyond  the  toe  .of  the  mountain  and  in  the 
valleys  the  sinking  always  encounters  plenty  of  water,  but  in 
the  two  exceptions  noted  the  location  of  the  shaft  in  each  case 
was  back  of  the  toe  of  the  mountains  and  away  from  the 
valleys.  The  quantity  of  water  in  the  wet  workings,  varies 
from  200  to  1100  gals,  per  min.  The  heavy  charge  of  high 
explosive  necessary  to  break  the  dense  sandstone,  coupled  with 
the  large  volume  of  water,  gives  the  worker  constant  pump 
trouble;  moving  the  pumps  during  the  blasting  is  out  of  the 
question,  because  the  water  would  flow  in  so  fast  that  the 


188  COAL  MINING  COSTS 

pumps  could  not  be  replaced  for  service.  Having  this  constant 
pump  trouble  makes  progress  very  slow,  reducing  it  to  10  ft. 
per  month  in  several  instances.  The  following  is  a  list  of 
some  of  the  shafts  and  amount  of  water  encountered: 

The  Middlestate  Coal  Co.  started  an  old  shaft,  abandoned 
90  ft.  down,  which  flowed  1100  gals,  of  water  per  minute,  and 
they  were  about  18  months  getting  down  the  additional  90  ft. 

The  Pocahontas  Collieries'  16  X  32  ft.  shaft,  at  Boissvain, 
was  11  months  in  sinking  200  ft.,  hindered  as  it  was  by  500 
gals,  of  water  per  minute. 

The  most  expensive  shaft  proposition  was  the  shaft  for  the 
Jed  Coal  &  Coke  Co.  about  2  miles  from  Welch.  This  property's 
main  shaft  encountered  1100  gals,  per  min.  while  its  air  shaft 
met  500  gals.  The  coal  expense  for  power  has  been  as  high  as 
$1500  per  month  with  coal  at  about  $1.25  per  ton. 

Sinking  costs. — The  Nokomis  Coal  Co.  sunk  a  shaft  through 
the  soft  Illinois  shales  in  1913,  the  cost  figures  on  which  are 
of  interest.  The  shaft  was  631  ft.  deep,  timber  lined  and  17  ft. 
5  in.  X  11  ft.  5  in.  inside  the  timbers,  with  an  airshaft  of  the 
same  dimensions  500  ft.  distant.  Eight  hand-feed  hammer 
drills,  weighing  40  Ib.  each  were  used  on  the  work  and  com- 
pressed air  was  furnished  by  a  9  X  10  X  12-in.  compressor 
supplying  174  cu.  ft.  of  free  air  per  minute  at  a  terminal  pres- 
sure of  100  Ib.  per  square  inch. 

After  passing  through  the  upper  capping  of  hard  rock, 
shale  of  various  degrees  of  hardness  was  encountered  with  an 
occasional  layer  of  limestone  10  or  12  ft.  thick.  The  rock  at 
times  consisted  of  slaty  bands,  sandy  shale  or  soft  gray 
material,  more  like  indurated  clay  than  shale. 

In  sinking  through  limestone,  from  28  to  32  holes  constituted 
a  round,  the  corner  holes  being  bottomed  2  in.  outside  the  line 
of  curbing.  As  the  shafts  were  timbered  throughout,  the  break 
lines  were  12  ft.  5  in.  by  18  ft.  5  in.  The  holes  were  4y2  ft. 
deep  and  were  connected  and  fired  with  an  electric  battery 
in  the  ordinary  manner.  Three  8-hr,  shifts  were  worked,  a 
round  being  drilled,  blasted  and  mucked  out  on  each  shift. 
Four  drillers,  four  muckers  and  a  shift  leader  or  boss  con- 
stituted a  sinking  crew. 

It  was  found  that  in  trying  to  bore  holes  in  the  soit  shale 
with  these  hammer  drills,  the  tool  cut  so  rapidly  as  to  choke 


SHAFT  SINKING  189 

the  passage  in  the  bit  with  muck,  stopping  the  flow  of  exhaust 
air  and  preventing  proper  cleaning  of  the  drill  holes.  Hand 
drilling  was  temporarily  substituted,  but  on  adopting  certain 
recommendations  of  the  manufacturers,  the  hammer  drills 
worked  satisfactorily,  and  their  use  was  resumed  with  a  marked 
increase  in  speed  over  the  hand  work. 

These  changes  gave  the  following  results  in  soft  shale,  as 
compared  with  hand  drilling.  Hand  drillers  using  21/4-in.  steel 
worked  at  the  rate  of  4y2  ft.  per  man  per  hour,  or  four  men 
put  in  a  round  of  54  ft.  in  3  hr.  With  the  air  drill,  four  men 
drilled  18.9  ft.  per  man  per  hour,  or  a  round  of  54  ft.  in  45  min. 
thus  accomplishing  a  saving  of  2  hr.  and  15  min.  per  54-ft 
round. 

Owing  to  the  variation  in  the  time  required  for  the  shoot- 
ing and  mucking,  this  increase  in  drilling  speed  meant  an 
increase  in  depth  per  day  of  1  ft.  or  4%  ft.  per  24  hr.,  with 
hand  drills.  Sinkers,  including  drillers  and  muckers,  were  paid 
$3.39  per  8-hr,  shift,  the  shiftboss  receiving  $4,  making  a  total 
labor  cost  per  day  of  roughly  $93,  or  $26.50  per  foot  by  hand 
drilling.  The  saving  by  using  the  hammer  drills  therefore, 
amounted  to  one  foot  in  each  shaft  or  $53  per  24  hr.  for  both 
shafts. 

Interesting  data  on  the  comparative  cost  of  sinking  a  ver- 
tical shaft  and  an  inclined  slope  to  accomplish  the  same  pur- 
pose were  disclosed  in  a  request  for  tentative  bids  made  to  two 
leading  contracting  companies  early  in  1921.  The  conditions 
under  which  the  work  was  to  be  done  were  as  follows : 

The  shaft  to  be  in  Ohio,  5  miles  from  present  railroad. 
Fuel,  supplies,  and  equipment  to  be  hauled  in  by  the  com- 
pany. Company  will  furnish  all  timber.  Strata  shale  and  sand- 
stones. Plenty  of  water  available.  Company  will  provide 
housing  accommodations.  Contractors  to  furnish  all  equipment 
except,  ties,  rails,  and  such  supplies  as  can  be  used  in  perma- 
nent mine  equipment.  Waste  disposal  within  radius  of  500  ft. 
Assume  an  average  amount  of  water  in  sinking. 

First  condition. — Shaft  three  compartments  10  X  26  ft.  in 
clear  170  ft.  deep  to  coal.  Timbered  with  8  X  10  in.  sets  on 
5  ft.  centers  with  2  in.  lagging. 

Second  condition. — Single  track  slope  on  30  deg.  pitch  5  ft. 
clear  above  top  of  rail  by  7  ft.  wide,  340  ft.  long,  timbered 


190  COAL  MINING  COSTS 

with  6  X  8  in.  crossbars  and  posts  with  sets  spaced  5  ft.  centers 
and  lagged  with  2  in.  planks  for  170  ft. 

Airshaft  within  500  ft.  of  above  slope  mouth  10  X  10  ft. 
and  170  ft.  deep  timbered  and  lagged  for  100  ft.  with  8  X  10  in. 
sets  on  5  ft.  centers. 

Third  condition. — Same  as  first  paragraph  of  second  con- 
dition except  duplicate  slopes  with  35  ft.  horizontal  pillar  be- 
tween. No  air  shaft. 

Fourth  condition. — Same  as  second  condition  except  double 
track  slope  12  ft.  wide.  Quote  both  with  and  without  air- 
shaft. 

The  first  company  submitted  bids  as  follows: 

The  price  of  the  shaft  in  the  first  condition  will  be  $200 
per  vertical  foot  timbered. 

Under  the  second  condition,  the  price  of  the  single  track 
slope  timbered  will  be  in  the  neighborhood  of  $100  per  lineal 
foot.  An  airshaft  under  this  condition  would  cost  in  the 
neighborhood  of  $140  per  vertical  foot  timbered  and  slopes 
under  the  third  condition  would  be  $100  each. 

Under  the  fourth  condition,  the  double  track  slope  would 
cost  approximately  $125  per  lineal  foot  sunk  at  the  same  time 
as  an  airshaft  is  being  sunk.  If  no  other  opening  is  sunk  at 
the  same  time,  that  is,  with  the  same  main  plant  and  the  same 
overhead,  the  slope  would  cost  $140  per  lineal  foot. 

The  price  for  cement  used  as  grout  will  run  in  the  neighbor- 
hood of  $12  per  barrel. 

The  bids  of  the  second  company,  which  were  gross,  were 
as  follows: 

First  condition. — Three  compartment  shaft,  10  X  26  ft.  in 
the  clear,  170  ft.  deep,  to  cost  $45,500. 

Second  Condition. — Single  track  slope,  5  ft.  clear  above  top 
rail,  7  ft.  wide,  340  ft.  long  and  airshaft  10  X  10  ft.,  to  cost 
$42,500. 

Third  condition. — Same  as  second  condition,  except  duplicate 
slopes,  no  airshaft,  to  cost  $42,400. 

Fourth  condition. — Same  as  second  condition,  except  double 
track  slope,  12  ft.  wide  with  airshaft,  $47,000 ;  without  airshaft 
$33,250. 

There  are  certain  items  entering  into  the  cost  of  such  a 
short  job  as  this  which  run  the  cost  up  considerably;  for 


SHAFT  SINKING 


191 


instance,  the  item  of  transportation  in  the  first  proposition, 
would  make  a  cost  of  about  $8  per  foot  of  shaft. 

Sinking  costs  given  below  were  fairly  representative  for 
the  different  methods  of  work  in  1909 : 

Shaft  excavated  14  X  20y2  ft.  through  6  ft.  of  soil  and  14  ft. 
of  quicksand,  not  very  wet.  Sides  supported  by  2-in.  oak 
sheeting  driven  by  mauls  and  braced  by  five  sets  of  10  X  12  in. 
timber : 


Per  Foot 

Per  Cubic  Yard 

Labor  

$27  25 

$2  57 

Lumber,  6600  ft.  B.M.  at  $30  

9  90 

93 

Erection  of  derrick,  etc  

3  00 

29 

Superintendence  

3  00 

29 

Sundry  

2  00 

18 

Coal  and  pumping  

5  00 

47 

Total 

$50  15 

$4  73 

Shaft  excavated  12  X  20  ft.  3  in.  through  45  ft.  of  clay  and 
gravel.  Sides  supported  by  sets  of  10  X  10  in.  pine  timber 
spaced  4%  ft.  centers  and  hung  from  top.  IV^-ni-  lagging: 


Per  Foot 

Per  Cubic  Yard 

Labor  

$19  50 

$2  17 

Lumber,  240  ft.  per  foot  at  $25  
Bolts,  15  Ib.  per  foot  at  $0.03  
Erection  of  head-frame,  etc  

6.00 
.45 
2  00 

.66 
.05 
22 

Superintendence  .... 

2  00 

22 

Power  . 

1  50 

17 

Sundry 

1  00 

11 

Total  

$32  45 

$3  60 

Shaft  excavated  15  X  37  ft.  through  21  ft.  of  dry  sand. 
Sides  supported  by  interlocking  steel  sheet  piling  driven  with 
steam  hammer  and  braced  with  sets  of  8  X  10  in-  timber : 


192 


COAL  MINING  COSTS 


Labor  Costs  Only 

Per  Foot 

Per  Cubic  Yard 

Driving  sheeting  

$6.55 

$0  32 

Removing  sheeting  
Timbering 

1.85 

2  05 

.09 
10 

Fixcavation  

8.20 

40 

Total  .  . 

$18.65 

$0.91 

Caisson  26  ft.  outside  diameter,  21  ft.  inside  diameter,  sunk 
through  56  ft.  semiliquid  mud  and  boulders: 


Per  Foot 

Per  Cubic  Yard 
Excavation 

(Materials  

$27  .  00 

$1.35 

Labor 

7  00 

35 

Forms  and  shoe  .  

23.00 

1.15 

Sinking  caisson            

38  00 

1.90 

Plant  erection 

3  00 

15 

Superintendence  

5.00 

.25 

Sundry 

5  00 

25 

Coal  and  power  

6.00 

.30 

Total  

$114.00 

$5.70 

The  H.  C.  Frick  Coke  Co.  put  down  two  shafts  nearly 
600  ft.  deep  near  Brownsville,  Pa.,  about  1909  that  presented 
some  interesting  cost  data.  The  main  shaft  is  elliptical  in  shape 
and  13  X  28  ft.  in  the  clear,  the  inside  circumference  being 
69  ft.  with  a  clear  opening  of  310  sq.  ft.  The  airshaft  is  the 
same  shape,  14  X  34  ft.  on  its  main  center  lines,  measures  81  ft. 
around  the  circumference  and  has  a  clear  opening  of  390 
sq.  ft. 

All  concrete  for  lining  the  shafts  is  composed  of  one  part 
Portland  cement;  two  parts  clean,  sharp,  river  sand,  and 
five  parts  of  stone  crushed  to  pass  through  a  IVk-ni-  ring. 
About  50  per  cent  of  the  stone  used  for  concrete  was  obtained 
from  the  materials  excavated  from  the  shafts.  About  30  per 


SHAFT  SINKING  193 

cent  was  shipped  in,  crushed  ready  for  use,  while  about  20 
per  cent  was  obtained  from  a  quarry  on  the  grounds. 

The  average  amount  of  concrete  to  each  batch  was  %  cu.  yd. 
Through  solid  strata  the  proportion  of  the  mixture  is  one  part 
cement,  two  parts  sand,  and  five  parts  crushed  stone,  while 
through  softer  strata  the  amount  of  cement  was  increased  50 
per  cent,  the  other  ingredients  remaining  the  same.  The  mini- 
mum thickness  of  the  concrete  lining  wall  is  12  in.  In  soft 
strata  the  concrete  is  as  much  as  33  in.  in  thickness,  as  no 
voids  were  left  between  the  rock  and  lining,  all  such  being 
carefully  filled  with  concrete. 

During  the  excavating  of  the  shafts,  blasting  was  done 
within  10  ft.  of  the  concrete,  but  at  no  time  did  this  blasting 
have  any  appreciable  effect  upon  the  concrete  lining.  Sixty- 
per-cent  gelatine  was  used  in  the  blasting,  and  the  maximum 
charges  were  from  150  to  200  Ib.  per  shot.  The  holes  in  the 
"cut,"  or  sump,  were  usually  drilled  to  a  depth  of  12  ft.,  while 
those  in  the  side  rounds  were  10  ft. 

The  sinking  and  concreting  were  carried  on  separately. 
The  work  was  pushed  continuously  from  12  o'clock  Sunday 
nights  until  12  o'clock  the  following  Saturday  nights.  No 
Sunday  work  was  done,  except  in  rare  cases,  other  than  that 
necessary  for  pump  operation.  All  men  employed  in  the  shafts 
worked  8-hr,  shifts.  No  forms  were  removed  under  72  hr., 
allowing  ample  time  for  the  concrete  to  set.  The  average 
time  for  completing  one  50-ft.  section  of  the  concrete  lining 
wall  occupied  6  days,  during  which  operation  the  services  of 
9  men  at  the  top  and  4  men  on  the  bottom  platform  were 
required.  In  sinking,  5  men  were  used  on  top  and  an  average 
of  12  men  on  the  bottom.  An  auxiliary  hoisting  engine  was 
used  in  one  compartment  for  the  handling  of  the  steel  forms; 
the  operation  of  placing  and  removing  these  forms  required 
the  services  of  4  men  at  the  top  of  the  shaft  and  6  men  on  the 
bottom. 

A  special  feature  of  the  work  was  the  construction  of  a 
4-in  curtain  wall  in  the  airshaft.  The  reinforcing  in  this  wall 
consisted  of  %-in.  round  steel  frames,  5  ft.  high  by  13  ft.  6  in. 
long,  stiffened  with  two  additional  %-in.  round  steel  placed 
equal  distance  from  the  ends.  Around  the  bars  forming  the 


194  COAL  MINING  COSTS 

frame,  No.  10  gauge  wire  netting  was  laced,  this  netting  having 
a  2-in.  mesh.  The  concrete  used  in  this  curtain  wall  was  made 
of  one  part  of  cement,  and  two  parts  of  clean  coarse  river  sand. 

In  the  excavating  of  the  shafts  a  record  was  kept  of  the 
muck  taken  out  and  of  the  materials  entering  into  the  work. 
From  the  main  hoisting  shaft  21,168  buckets  of  muck  were 
taken,  while  the  ventilating  shaft  gave  28,442  buckets,  approxi- 
mating 50,000  cu.  yd.  of  loose  excavation  through  earth,  fire- 
clay, shale,  slate,  sandstone,  limestone,  and  coal.  The  increase 
of  the  actual  muck  excavated,  after  the  blasting,  over  the 
calculated  yardage  in  the  solid  shows  135.2  per  cent  for  all  the 
materials  passed  through,  or  35.2  per  cent  more  than  double 
the  amount  in  the  solid. 

There  were  used  in  the  construction  of  the  concrete  lining 
in  the  shafts,  8217  barrels  of  cement,  4410  tons  crushed  stone 
and  2528  tons  sand,  made  up  into  12,951  batches  of  concrete, 
containing  approximately  6500  cu.  yd. 

Exclusive  of  the  archways  at  the  bottom  landings,  539  ver- 
tical feet  of  concrete  lining  wall  was  placed  in  the  hoisting 
shaft.  The  time  to  complete  the  same  covered  a  period  of  46 
weeks  from  the  start  of  the  work  of  excavating,  showing  a 
progress  of  48  ft.  per  month. 

Omitting  archways  at  the  bottom  landing,  561  vertical 
feet  of  concrete  lining  was  placed  in  the  ventilating  shaft. 
The  time  consumed  in  completing  same  covered  a  period  of 
62  weeks  from  the  commencement  of  the  work  of  excavating; 
a  progress  of  37.2  ft.  per  month,  for  the  sinking,  timbering, 
and  placing  of  the  concrete  lining.  Taking  into  consideration 
that  this  shaft  has  an  area  of  80  sq.  ft.  more  than  the  hoisting 
shaft  the  progress  of  the  work  averaged  about  the  same  as 
at  the  other  shaft. 

The  first  cost  of  concrete-lined  shafts  over  that  of  timber 
lined  is  about  one-third  more,  which  amount  would  probably 
be  spent  upon  the  first  renewal  of  timbers  and  from  the  aver- 
age run  of  timber  now  on  the  market  this  would  likely  be 
necessary  in  about  10  yr.  In  a  concrete-lined  shaft,  the  lining 
being  indestructible,  the  only  renewals  required  would  be 
replacing  of  the  buntons  and  guide  rails  from  time  to  time. 


SHAFT  SINKING 


195 


COMPARATIVE  QUANTITIES  IN  THE  CONCRETE  LINING  FOR  THE  VERTICAL 

FOOT  OF  SHAFT 

HOISTING  SHAFT 


Calculated  Yardage 

Actual  Yardage 
of  Material 
Placed  in  Work 

Increase  of  Actual 
Yardage  over  the 
Yardage  Calculated 

Thickness,  Inches 

Cubic  Yards 

Cubic  Yards 

Percentage 

12 
15 

18 
24 

2.66 
3.36 
4.08 
5.56 

4.6 
5.0 
6.4 
9.0 

73 
49 
57 
62 

VENTILATING  SHAFT 


12 

3.10 

5.40 

74 

15 

3.92 

5.62 

43 

18 

4.74 

5.84 

23 

24 

6.45 

9.50                             47 

The  following  summary  gives  the  principal  figures  in  regard 
to  the  work  on  both  shafts: 

HOIST  SHAFT 

Total  depth,  feet 565 

Total  number  of  weeks  worked 46 

Average  thickness  of  concrete  lining,  inches 16 

Average  number  of  batches  per  5-ft.  form 56 

Average  cubic  yards  concrete  per  5-ft.  form 28 

Average  depth  of  lining  placed  per  week,  feet 12 . 3 

Average  depth  of  sinking  and  lining  placed  per  month,  feet 48 

VENTILATING  SHAFT 

Total  depth,  feet 591 .2 

Total  number  of  weeks  worked 62 

Average  depth  sunk  per  week,  feet 9.5 

Average  depth  of  lining  placed  per  week,  feet 9.1 

Average  depth  of  sinking  and  lining  placed  per  month,  feet 37 . 2 

Size  of  shaft,  inside  concrete  lining  wall 14  ft.  X 34  ft. 

Area,  square  feet .  390 

Average  thickness  of  concrete  lining,  inches 17 

Average  number  of  batches  per  5-ft.  panel 66 

Average  cubic  yards  of  concrete  per  5-ft.  panel 33 


196  COAL  MINING  COSTS 

Some  interesting  cost  figures  were  obtained  in  the  sinking 
of  a  570-ft.  inclined  metal  mining  shaft  in  the  Poverty  Gulch 
region  of  Colorado  about  1910.  The  shaft  was  12  X  7.5  ft. 
in  the  clear  and  after  it  had  been  sunk  a  short  distance  two 
skips  were  added  to  the  sinking  equipment  which  weighed 
550  Ib.  each,  held  10  cu.  ft.  and  cost  $70  a  piece,  including  a 
water  valve  in  the  bottom  which  cost  $10.  The  sinking  equip- 
ment consisted  of  the  following: 

Six  Little  Giant  drills,  2|-in.  diameter. $900.00 

Six  columns  and  arms  up  to  8  ft 240 . 00 

Six  sets  drill  steel,  each  15.4  Ib.,  at  $14.40 86 . 40 

Buffalo  exhaust  fan,  diameter  outlet  24£  in.,  price  with  bed  and 

countershaft 700 . 00 

460  ft.  of  24-in.  pipe,  at  $54  per  100  ft 248.40 

1200  ft.  of  12-in.  pipe,  at  $27  per  100  ft 324 . 00 

Ten  Leyner  No.  5  stoppers,  $135 1350.00 

Drill  steel,  10  sets,  at  $15 150 . 00 

Sixteen  ore  cars,  at  $45 720 . 00 

Compressor,  motor  and  pipe,  freight 2689 . 14 

Freight  on  19,553  Ib.,  at  $0.55  per  100 107.54 


Total $7515.48 

In  the  development  work  two  drills  were  required.  The 
drills  are  2%  in-  in  diameter  and  use  air  at  90  Ib.  pressure  per 
square  inch  at  drill.  According  to  the  catalog  specifications, 
each  drill  will  need  67.2  cu.  ft.  of  air  per  minute.  The  factor 
to  determine  a  compressor  capacity  for  four  drills  at  10,000  ft. 
altitude  is  4.49;  hence,  301.5  cu.  ft.  of  air  per  minute  will  be 
required,  but  deducting  5  per  cent  for  leakage  and  allowing 
the  compressor  a  volumetric  efficiency  of  80  per  cent,  the  total 
air  required  is  nearly  400  cu.  ft.  per  minute. 

E.  A.  Rix  allows  20  hp.  for  every  100  cu.  ft.  of  cylinder 
displacement,  to  compress  air  to  90  or  95  Ib.  gauge  pressure 
at  sea  level.  Although  20  hp.  is  higher  than  the  value  given 
by  Peele  for  the  theoretical  horsepower  required,  and  figuring 
efficiency  the  figures  would  then  be  below  20;  but,  since  com- 
pressors are  usually  purchased  for  excess  power  to  supply 
possible  additional  uses,  and  the  use  of  20  hp.  would  only  add  a 
small  percentage  on  the  safe  side,  the  power  necessary  for  four 
drills  is  taken  at  80  hp.,  and  it  was  decided  to  purchase  a  two- 


SHAFT  SINKING  197 

stage  air  compressor  18  X  11  in.  diameter,  12-in.  stroke, 
125  r.p.m.,  with  a  capacity  of  440  cu.  ft.  at  10,000  ft.  elevation. 

This  sized  compressor  gives  a  reserve  of  21  per  cent,  and 
costs  with  freight  from  Denver  to  Cripple  Creek  $1903.'  The 
motor  for  the  compressor  is  80  hp.;  900  r.p.m.,  440  volts,  and 
costs  delivered  $710.84. 

The  following  calculations  give  the  power  consumed  during 
development  by  two  piston  drills:  The  catalog  multiplier  is 
2.39  and  as  each  drill  will  need  67.2  cu.  ft.  of  air,  67.2  X  2.39 
=  159.  Allowing  for  air  loss  in  pipe  line  and  efficiency  of 
compressor,  210  cu.  ft.  are  necessary,  or  42  hp.  per  minute. 

Each  stope  drill  requires  25  cu.  ft.  of  air  per  minute  and  the 
factor  for  seven  drills  is  7.55 ;  therefore,  25  X  7  =  189  cu.  ft., 
to  which  63  cu.  ft.  is  added  to  allow  for  loss  and  efficiency, 
and  this  is  equivalent  to  50.4  hp. 

The  power  for  two  stope  drills  is  25  X  2.5  (multiplier) 
=  62.5  cu.  ft.  of  air,  and  if  to  this  be  added  20.7  cu.  ft.  for 
pipe  loss  and  efficiency,  the  power  required  is  16.68  hp.  The 
diameter  of  the  pipe  needed  for  carrying  air  800  ft.  is  3  in. 
and  will  cost  at  Cripple  Creek  $75.30.  The  total  cost  of  com- 
pressor, motor,  and  pipe  is  $2,689.14. 

In  sinking,  18,  4-ft.  holes  were  used.  Drilling  was  done 
at  the  rate  of  39  ft.  in  8  hr.  per  drill  or  72  ft.  in  7.4  hr.  The 
rate  of  advance  was  3  ft.  per  round  which  equals  3  X  12  X  7.5 
=  270  cu.  ft.  solid  material  which  divided  by  12.4  gives  21.8 
tons  per  round  and  multiplying  this  by  21.5  gives  469  cu.  ft. 
or  17.4  cu.  yd.  of  loose  material.  Estimating  the  cost  of  muck- 
ing on  the  basis  of  1.2  cu.  yd.  per  man  per  hour  it  will  take 
two  men  7%  hr.  to  clean  up  the  rock  after  each  round,  so 
that  allowing  %  hr.  for  delays,  changing  buckets,  etc.  it  will 
be  seen  that  one  round  can  be  drilled,  fired  and  mucked  in 
16  hr.,  which  eliminating  other  delays  would  be  equal  to  a 
progress  of  90  ft.  per  month. 

For  a  4-ft.  hole  it  was  found  that  six  sticks  of  40  per  cent 
dynamite  were  required  each  weighing  0.6  Ib.  or  3.6  Ib.  per  hole 
so  that  for  the  18  holes  64.8  Ib.  were  required. 

The  shaft  was  sunk  570  ft.  The  time  required  to  sink  the 
shaft  was: 

5-77 — -    -7  =  190  rounds,  or  days. 

3  ft.  per  round 


198  COAL  MINING  COSTS 

The  detailed  cost  of  sinking  570  ft.  of  a  90-sq.  ft.  inclined 
shaft  is  as  follows: 

Two  machine  men,  190  shifts,  at  $4.50 $1,710.00 

Two  muckers  (also  top  men)  190  shifts  at  $3 1,140 . 00 

Two  hoistmen,  190  shifts  at  $4.50 1,710.00 

One  blacksmith,  190  shifts  at  $4.50 855.00 

One  blacksmith  helper,  190  shifts  at  $4 760 . 00 

One  foreman,  190  shifts  at  $4 . 50 855 . 00 

One  superintendent,  at  $175  per  month,  6^  months 1,108. 35 

One  timberman,  190  shifts,  at  $3.50 665 .00 

Powder,  12,312  lb.,  at  $1 .27* 1,563 . 62 

Fuse,  190  rounds,  7-ft.  lengths,  23,940  ft.,  at  $0 . 0035* 83 . 79 

Caps,  3420,  at  $0.007* 23 .94 

Depreciation  on  steel 14 . 40 

Operation  compressor  plant  (power),  336  hp.  hours  per  day: 

Installation  charge $20 . 50 

40,000  kw.  hr.,  at  $0.013 520.00 

7700  kw.  hr.,  at  $0.005 38.50  579.00 

Timber,  95,440  ft.,  at  $20  per  thousand 1,908 . 80 

Electric  power,  for  hoist,  average  6.23  hp.  hr.  per  hour,  49.84 

hp.  hr.  per  day,  9470  kw.  hr.  (190  days)  at  $0 . 013 123 . 11 

Coal  for  blacksmith,  28.75  tons,  at  $20.75 243 .20 

Candles,  950,  at  $0. 0145 13 . 78 

Rails  (30-lb.),  570  ft.,  5.08  tons,  at  $50 254.00 


Cost  for  570  ft $14,207.55 

Or  cost  per  ft $24.93 

*  These  low  costs  of  powder  are  those  of  the  Portland  Gold  Mining  Co.  and  include 
freight  and  unloading  charges. 

The  following  costs  of  sinking  a  mine  shaft  through  ande- 
site  at  the  Esperanza  Mine  at  El  Oro,  Mexico,  are  given  by 
W.  E.  Hindry  in  the  Mining  and  Scientific  Press  in  1910.  The 
shaft  was  a  three-compartment  vertical  shaft,  having  two 
5X5  ft.  hoisting  compartments  and  a  5X7  ft.  pump  and 
ladderway.  The  timbering  was  10  X  10  in.  with  2-in.  lagging ; 
sills  5  ft.  center  to  center,  and  6  posts  per  set.  The  total  depth 
of  the  shaft  was  679  ft.,  of  which  101  ft.  were  sunk  by  wind- 
lass and  hand  work,  and  578  ft.  by  steam  hoist  and  machine 
drills.  The  work  was  done  in  1899  and  the  prices  of  materials 
and  wages  were  as  follows : 


SHAFT  SINKING 


199 


Materials 


Prices 


Timber  per  M  ft.  B.M $13.58 

Wood  per  cord 3 . 15 

Coal  per  ton 7.27 

Powder,  60  per  cent,  per  pound 0. 14f 

Fuse  per  foot 0.0055 

Caps,  each 0.0058 

Candles,  each 0.0194 

Labor 

Superintendent,  per  24  hr :'.'„..  '  $4.850 

Shaft  men,  foreign,  per  8-hr,  shift 3 . 220 

Shaft  men,  native,  per  8-hr,  shift 0.528 

Top  men,  per  8-hr,  shift 0.422 

Fireman,  per  8-hr,  shift 0.485 

Hoistmen,  per  8-hr,  shift 0.970 

Blacksmiths,  per  8-hr,  shift 1 .455 

The  cost  of  excavation  was  as  follows : 

Labor  Per  Linear  Foot 

Superintendence $2 . 529 

Shaftmen,  foreign 3 . 510 

Shaftmen,  native 7.043 

Top  men 0. 678 

Blacksmiths 0.718 

Firemen 0.317 

Hoistmen 0. 894 

Miscellaneous 0.936 

Total $16.525 

Materials 

Timber $3.961 

Wood,  fuel 3.781 

Coal 0. 179 

Powder 2.853 

Fuse 0 . 014 

Caps 0.294 

Candles 0.223 

Oil,  grease,  etc 0.025 

Miscellaneous 0 . 051 

Total $11.381 

Grand  total..  $27.906 


200 


COAL  MINING  COSTS 


The  above  costs  are  converted  from  Mexican  money  assum- 
ing the  peso  to  have  a  value  of  48^c. 

An  exhaustive  study  of  shaft  sinking  costs  in  the  Michigan 
region  was  prepared  about  1910.  The  figures  and  computa- 
tions were  made  upon  the  assumption  that  the  shaft  would  be 
6  X  16  ft.  within  timbers  and  reach  a  vertical  depth  of  1000  ft. 
For  a  vertical  shaft,  its  total  length  would  be  1000  ft. ;  if  in- 
clined at  an  angle  of  45  deg.  to  follow  the  dip  of  the  formation, 
its  total  length  would  be  1400  ft.  to  reach  a  total  vertical 
depth  of  1000  ft.  A  contract  price  of  $40  per  foot  for  sinking 
would  apply  only  if  the  shaft  was  put  down  in  the  jasper 
formation.  If  diorite  was  encountered  the  contract  price  for 
sinking  alone  would  be  between  $50  and  $55  per  foot,  while 
all  other  items  would  remain  the  same. 

A  slight  difference  in  cost  of  timber  would  appear  in  the 
amount  of  timber  used  in  a  vertical  or  inclined  shaft,  as  the 
latter  requires  10  X  10  in.  stringers,  while  the  former  would 
take  6X8  in.  skip  runners,  but  this  has  not  been  taken  into 
account. 

The  maximum  flow  of  water  to  be  handled  (800  gal.  per 
min.)  is  probably  somewhat  high. 

DETAILS  OP  SINKING  CONTRACT,  SHAFT,  ETC. 


Maximum 


Minimum 


Contract  price  for  sinking,  $40  per  foot,  includes 

drilling,  blasting,  powder,  caps,  fuse,  etc. 

For  a  vertical  shaft,  1000  ft 

For  an  inclined  shaft,  1400  ft 

Computing  all  work  for  a  shaft  6X16  ft.  within 

timbers,    three    compartments,    shaft    sets,    5-ft. 

centers;  and  assuming  40  ft.  per  month  as  average 

sinking. 

1000  ft.  would  take  25  months,  allow  26  months. 
1400  ft.  would  take  35  months,  allow  36  months. 
Allowing  25  working  days  per  month,  300  per  year 
1000  ft.  of  sinking  would  take  650  days. 
1400  ft.  of  sinking  would  take  900  days. 
Sinking  contract  would  be  worked  on  three  8-hr. 

shifts.     All  other  labor,  one  or  two  10-hr,  shifts. 
Cutting  pump  station  10X15X20  ft.  at  a  depth  of 

500ft 

A  maximum  flow  of  800  gal.  per  min.  when  bottom  of 

shaft  is  approached  is  used  as  the  basis  of  pumping 

expenses. 

Total..  


$56,000 


$40,000 


350 


350 


$56,350 


$40,350 


SHAFT  SINKING 
DETAILS  OF  LABOR 


201 


Maximum 


Minimum 


Blacksmith,  at  $2.25  per  day;    helper  at  $1.65  per 

day;  $3 . 90  per  day  for  900  days 

$3.90  per  day  for  650  days 

Two  landers  at  $1 . 70  per  day  for  900  days 

Two  landers  at  $1 . 70  per  day  for  650  days . 

Two  timbermen  at  $1.75  per  day.  Assuming  one 

set  of  timber  can  be  cut  and  framed  in  1  day  by 

two  men. 

1400  ft.  of  shaft,  280  sets,  280  days 

1000  ft.  of  shaft,  200  sets,  200  days 

Two  brakemen  at  $2 . 20  per  day,  for  900  days 

Two  brakemen  at  $2 . 20  per  day,  for  650  days 

Two  firemen  at  $1 . 70  per  day  for  900  days 

Two  firemen  at  $1 . 70  per  day  for  650  days 

Allow  one-fourth  of  mining  captain's  time,  $25  per 

month,  for  36  months 

Allow  one-fourth  of  mining  captain's  time,  $25  per 

month,  for  26  months 

Surveyor  and  helpers,  allow  $20  per  month,  for  36 

months 

Surveyor  and  helpers,  allow  $20  per  month,  for  26 

months . . 


$3,510 
'3,060 


980 
3,960 


3,060 
900 


720 


$2,535 
2,210 

700 
2,860 
2,210 

650 
520 


Total 


$16,190        $11,685 


DETAILS  OF  TIMBER 


Board 
Measure, 
Feet 

Maximum 

Minimum 

2  Plates,  12X12  in.XlS  ft.  contain  
2  End  pieces,  12X12  in.  X6  ft.  contain.  . 
4  Corner  posts,  12X12  in.X4  ft.  contain 
2  Diyidings,  10X12  in.X6  ft.  4  in.  con- 
tain 

432 
144 
192 

126f 

4  Stringers,  10X10  in.  X5  ft.  contain.  .  .  . 
4  Center  posts,  10X10  in.X4ft,  contain. 
Sheathing,  3  in.  X  5  ft,  X  44  ft.  contain.  .  . 
Boards,  1  in.X5  ft.X6  ft.  4  in.  contain.  . 

166f 
133| 
660 
31f 

Total  amount  of  timber  for  1  set 

1  886  £ 

Total  amount  of  timber  in  280  sets  
Total  amount  of  timber  in  200  sets  

535,173 

377,267 

At  $14  per  thousand  for  hemlock  timber: 
for  280  sets  $7  492  42  allow 

$7500 

for  200  sets  $5  181  7*  allow 

$5  200 

Ladders   17  ^c  per  ft    1400  ft 

245 

Ladders   17^  c  per  ft    1000  ft 

175 

Total 

$7745 

$5  375 

202 


COAL  MINING  COSTS 


DETAILS  OF  RAILS,  PIPE,  TIE-RODS,  ETC. 


Maximum 

Minimum 

.45-lb.  rails  for  double  skip  road,  allow  1500  ft.,  mak- 
ing 6000  linear  ft.  or  2000  yds.,  45  tons  at  $25 

$1  125 

200  pairs  fish-plates,  at  21c  

42 

500  Ib.  rail  spikes,  at  3c  

15 

1120  1|  in.X6  ft.   6  in.  round  iron  tie-rods,  at  2c, 
$487  46  allow  

500 

2240  nuts  and  washers,  $278  .  60,  allow  

300 

All  of  the  above  computed  for  1400  ft.  of  shaft. 
For  a  vertical  shaft  of  1000  ft.  depth,  rails,  fish-plates 
and  spikes  would  not  be  used. 
Tie-rods,  nuts,  and  washers  for  1000  ft.  of  shaft,  allow  . 

$550 

1400  ft.  of  10-in.  water-column  pipe 

1  820 

1000  ft.  of  10-in.  water-column  pipe 

1  300 

Drill  steel  

40 

30 

2-in.  steam  pipe;  allow  250  ft.  in  excess  of  length  of 
shaft      For  1650  ft  

150 

For  1250  ft 

110 

4-in.  air  pipe;    allow  250  ft.  in  excess  of  length  of 
shaft.     For  1650  ft. 

750 

For  1250  ft  

575 

Allow  a  maximum  of  5  tons  of  coal  per  day  for  26 
months  

2280 

For  36  months  

3285 

3000  Ib.  60d.  spikes  at  lOc.  per  pound  for  1400  ft  ... 
200  Ib.  lOd.  nails  at  lOc.  per  pound,  for  1400  ft  .... 
2160  Ib.  60d.  spikes  at  lOc.  per  pound,  for  1000  ft  ... 
145  Ib.  lOd.  nails  at  lOc.  per  pound,  for  1000  ft  

300 
20 

216 
15 

Total  

$8  347 

$5  081 

DETAILS  OF  PLANT 


Maximum 


Minimum 


Shaft  house  and  pockets 

Engine  house 

Boiler  house 

Powder  house 

Coal  trestle 

Boilers  (4) 

Hoisting  engine 

Compressor 

Skips  (2) 

Two  No.  3  Rand  drills,  complete 

Auxiliary  pump  at  500-foot  depth 

Sinking  pump 

Temporary  equipment  at  start,  small  hoist,  bucket, 
rope,  tripod,  etc.,  allow 

Hoisting  cable,  If -in.  diameter 

Incidentals — Teaming,  pipe  fittings,  air  hose,  picks, 
shovels,  hammers,  wrenches,  timber  cutter's  tools, 
axes,    saws,    oil,    waste,    candles,    temporary   bell 
signal  system,  etc 

Total .  . 


$14,800 

9,625 

4,200 

250 

3,000 

11,660 

15,000 

11,500 

1,000 

375 

6,000 

1,000 

1,500 
900 


5,000 


$85,810 


$6,750 

5,000 

3,500 

100 

2,500 

10,500 

12,000 

10,625 

300 

375 

5,000 

900 

1,000 
700 


4,500 


$63,750 


SHAFT  SINKING 
RECAPITULATION 


203 


Inclined 
Shaft, 
1400  ft. 

Vertical 
Shaft, 
1000  ft. 

Sinking  contract  .... 

$56  350 

$40  350 

Blacksmithing  .  .  . 

3  510 

2  535 

Landers  

3  600 

2210 

Timber  cutters  

980 

700 

Brakemen  

3  960 

2  860 

Firemen  

3  060 

2  210 

Captain  and  surveyors  

1  620 

1  170 

Timber  and  ladders  

7745 

5  375 

Rails,  fish-plates,  spikes  

1  182 

Air  and  steam  pipes  and  water  column  . 

2720 

1  985 

Tie-rods,  nuts,  and  washers  

800 

555 

Nails  and  spikes  

320 

231 

Coal  

3285 

2280 

Drill  steel  

40 

30 

Total  

$88  632 

$62  491 

Total  plant  maximum $85,810 

Total  plant,  minimum 63,750 

Inclined  shaft  with  minimum  plant 152,382 

Inclined  shaft  with  maximum  plant 174,442 

Vertical  shaft  with  minimum  plant 126,241 

Vertical  shaft  with  maximum  plant 148,300 

In  spite  of  very  difficult  sinking  problems,  as  compared 
with  conditions  in  this  country,  shaft-sinking  costs  in  Europe 
have  been  substantially  less  than  in  this  country.  Shafts  in 
the  Taff  and  Rhonddha  valleys  in  England,  which  are  circular 
and  from  17  to  21  ft.  in  diameter  were,  about  1910,  sunk  at 
a  total  cost  of  $30  to  $50  per  foot  including  the  lining.  In 
the  north  of  England  the  shafts  are  somewhat  larger,  varying 
from  20  to  24  ft.  in  diameter  and  are  usually  lined  with 
steel  tubbing.  Some  excellent  speed  records  have  been  made 
at  these  shafts,  at  the  Sherwood  colliery  for  instance  a  shaft 
was  sunk  858  ft.  in  21  weeks,  an  average  of  40.8  ft.  per  week. 

In  Belgium  brick  lining  was  used  almost  exclusively  at  one 
time,  though  the  use  of  reinforced  concrete  is  becoming  more 
general.  In  1910  sinking  costs  there  were  about  $60  to  $75 
per  meter  and  the  lining  $5  to  $6  additional. 

It  is  frequently  necessary  to  put  down  small  prospect  shafts 
for  depths  up  to  100  ft.  and  the  approximate  cost  of  such 
equipment  as  of  1907  was  as  follows: 


204  COAL  MINING  COSTS 

One  25-hp.  vertical  boiler $300 

One  5 X5  in.  Bacon  type  of  hoist 350 

One  2f-in.  steam  drill 180 

One  7-ft.  bucket 30 

One  18-in.  sheave  and  bearings 20 

200  ft.  of  ^-in.  wire  rope 13 

Lumber  for  head-frame,  hauling,  and  labor 107 


Total $1000 

The  items  for  blacksmith  shop,  bunk  house,  cook  house, 
etc.,  must  usually  be  added,  but  this  amount  will,  of  course, 
depend  upon  the  size  of  the  prospect  and  if  they  are  needed. 
To  provide  a  moderate  equipment  and  to  allow  a  certain 
amount  of  working  capital  another  $1000  should  probably  be 
provided. 

Shaft  lining's. — An  interesting  example  of  the  costs  of  a 
concrete  lined  metal  mining  shaft  is  that  of  the  Brier  Hill 
shaft  at  Vulcan,  Mich.,  sunk  about  1909.  This  shaft  is  cir- 
cular, 14  ft.  in  diameter  and  850  ft.  deep.  Steel  sets  made  up 
of  8-in.  channels,  13%  Ib.  per  foot,  placed  on  edge  and  spaced 
10  ft.  8  in.  on  centers  were  used.  Between  the  sets  there  are 
studdles  of  steel  channels  to  which  the  wooden  runners  for 
the  cage  are  bolted.  The  ladders  are  built  of  steel  and  the 
ladderway  and  skip  compartment  are  lined  with  galvanized 
corrugated  sheet  steel. 

The  line  of  the  shaft  was  through  some  old  workings  so 
that  it  was  possible  to  make  preliminary  openings  throughout 
the  entire  length  from  these  different  levels ;  this  opening  was 
made  6  X  8  ft.  and  a  20  X  20-ft.  square  shaft  sunk  from  the 
surface  to  connect  with  this.  Concreting  was  started  at  a 
depth  of  79  ft.  from  the  surface. 

The  concrete  used  was  mixed  in  the  proportion  of  one 
cement,  three  sand  and  six  stone  and  the  average  thickness 
of  the  lining  was  18  in.  with  a  minimum  of  6  in.  Measure- 
ments of  the  actual  excavation,  taken  every  3  ft.,  showed  an 
average  thickness  of  the  lining  of  19  in.  which  was  equivalent 
to  a  little  less  than  3  cu.  yd.  of  concrete  to  a  vertical  foot  of 
shaft.  Forms  made  of  %-in.  sheet  steel  were  used,  there  being 
two  sets  of  forms,  each  5  ft.  4  in.  high  and  each  set  consisting 
of  four  segments. 


SHAFT  SINKING  205 

After  the  excavation  of  the  shaft  to  the  proper  size  was 
carried  downward  for  a  distance  sufficient  to  put  in  three  or 
four  sets  of  steel,  a  platform  or  "curb"  is  laid  on  the  outside 
edge  of  the  shaft  all  around  near  the  bottom  at  the  proper 
distance  from  the  concrete  above  to  allow  for  the  number  of 
sets  proposed.  One  round  of  forms  is  then  set  upon  this  plat- 
form. Empty  boxes  are  put  in  at  the  bottom  to  form  "chute 
holes"  for  putting  concrete  into  the  form  below  at  the  proper 
time  and  a  platform  of  3-in.  plank  is  laid  over  the  top  of  the 
form.  The  concrete  is  then  lowered  in  the  kibble,  dumped 
on  the  platform,  slushed  off  and  tamped  in  around  the  out- 
side of  the  forms  completely  filling  the  space  between  the  forms 
and  the  rock. 

A  set  of  steel  is  then  lowered,  laid  in  the  hitches  left  in 
the  concrete  and  cemented  in.  The  round  of  forms  is  lowered 
upon  this  new  set,  expanded  to  the  proper  size  and  the  second 
round  of  forms  is  set  up  and  bolted  to  the  first  so  that  this 
time  concrete  is  deposited  for  a  height  of  10  ft.  8  in.  This  can 
be  done  in  an  eight-hour  shift,  and  12  hr.  are  sufficient  for  the 
concrete  to  set  so  that  it  is  possible  to  concrete  10  ft.  8  in.  of 
shaft  and  put  in  the  steel  work  in  less  than  two  days.  As 
the  space  between  the  steel  sets  is  10  ft.  8  in.  and  in  starting 
only  half  this  amount  of  concrete  is  put  in  before  a  set  is  laid, 
the  work  is  joined  to  the  older  concrete  above  by  one  round 
of  forms  of  5  ft.  4  in.  This  is  filled  through  the  spaces  or 
"chute  holes"  left  by  the  empty  boxes  previously  mentioned. 

No  attempt  was  made  to  make  a  record  of  speed.  The  best 
work  was  done  in  September  and  October,  1909,  which  resulted 
in  the  excavating  (enlarging  the  original  opening)  concreting 
and  putting  in  the  steel  of  1382/3  ft.  of  shaft  or  an  average  of 
69V3  ft.  per  month.  At  that  time  the  men  were  working  two 
shifts  a  day  or  11  shifts  a  week  of  10  hr.  each.  Six  men  con- 
stituted the  regular  shift  in  the  shaft  either  for  excavating  or 
constructing.  In  addition  to  the  shaft  work  proper  three  sta- 
tions have  been  cut  out  and  concreted. 

In  keeping  the  record  of  the  cost,  the  shaft  has  been  divided 
into  three  parts:  from  surface  to  ledge,  62  ft.;  from  top  of 
ledge  to  7th  level,  549.5  ft.;  and  from  the  7th  level  down, 
ultimately,  about  238.5  ft. 


206 


COAL  MINING  COSTS 


COST  OF  CONSTRUCTION 


Surface 
to  Ledge, 
62ft. 

Ledge  to 
7th  Level, 
549.5  ft 

Preliminary  excavation  i  

$13  07 

$18  46 

Final  excavation 

18  19 

15  10 

Steel  shaft  frames        

7  90 

7  90 

Steel  forms 

0  83 

0  83 

Temporary  surface  structures  and  equipment  .  .  . 
Construction 

10.18 
56  29 

10.18 
25  26 

Estimated  charge  for  compressed  air 

1  00 

1  00 

Total  per  foot 

$107  46 

$78  73 

Estimated  salvage  on  shaft  timbers  
Estimated  salvage  on  temporary  surface  struc- 
tures and  equipment  

0.50 
2.95 

0.50 
2  95 

Net  total  per  foot  

$104.01 

$76.28 

The  item  "construction"  includes  the  handling  of  the  steel 
forms,  depositing  concrete,  and  setting  and  erecting  the  steel 
frames.  These  figures  include  estimates  of  power  and  every 
charge  except  for  general  management  and  engineering.  The 
cost  of  the  first  85  ft.  below  the  7th  level  was  $87.19  per  foot. 
The  cost  of  fhis  circular  concrete-lined  shaft  is  about  the  same 
as  a  rectangular  shaft  of  the  same  capacity  with  steel  fram- 
ing. The  small  additional  cost  over  a  rectangular  shaft  with 
timber  framing  is  abundantly  justified  by  the  increased  safety 
and  permanence. 

One  of  the  earliest  concrete  shaft  linings  in  this  country 
was  put  in  by  the  River  Coal  Co.  near  Bridgeport,  Pa.,  about 
1905.  The  shaft  lining  measures  23  ft.  on  its  major  axis  parallel 
to  the  railroad  tracks  and  15  ft.  on  the  minor  axis,  inside 
measurements,  the  thickness  of  its  concrete  walls  varying  with 
the  depth  below  the  surface.  The  following  table  gives  approxi- 
mately the  cost  of  the  construction  both  per  foot  of  depth  and 
per  cubic  yard: 


SHAFT  SINKING 


207 


Per 
Foot  of  Depth 

Per 
Cubic  Yard 

Stone                    

$5.90 

$1.00 

Sand                             

1  77 

0  30 

Cement 

19  18 

3  25 

Labor: 
Mixing                     

$3  83 

$0  65 

Placing 

3  40 

0  58 

Firemen  and  pumpmen  

2.19 

0.37 

940 

1   (\c\ 

Forms  : 
Lumber,  $13  per  thousand  

$1  83 

$0.31 

Making  $21  per  thousand.        

2  95 

0  50 

Placing 

4  82 

0  81 

9R1 

1      AO 

Platform  for  starting  upper  section 

0  92 

0  16 

Superintendence  

3.03 

0.51 

Plant                            

0  28 

0.05 

Oil 

0  21 

0  04 

Sundry             

1.06 

0.18 

Tools                                        

0  24 

0  04 

$51.62 

$8.75 

A  novel  shaft  lining  in  the  form  of  concrete  blocks  was 
used  in  a  shaft  in  Belgium  in  1912,  the  estimated  cost  of  which 
was  $8.65  per  running  foot  of  shaft.  This  lining  was  found 
to  be  equal  in  strength  to  a  32-in.  masonry  lining  the  cost  of 
which  would  have  been  $13.50  per  foot. 

The  shaft  was  133  ft.  deep  and  13  ft.  4  in.  in  diameter,  the 
excavation  being  about  17  ft.  in  diameter.  The  concrete  blocks 
were  30  in.  high  with  a  minimum  thickness  of  3.2  in.  and  14 
were  required  to  make  the  circumference  of  the  shaft.  They 
were  set  with  joints  staggered  and  the  successive  rings  of 
blocks  were  joined  by  12  X  0.6-in.  dowels  there  being  two  of 
these  to  each  block.  Concrete  with  additional  reinforcing  was 
filled  in  between  the  back  of  the  blocks  and  the  excavation. 

The  increased  cost  of  timber  and  inferior  product  being 
offered  in  recent  years,  has  tended  to  cause  the  use  of  other 
materials  for  shaft  linings,  especially  cement  which  has  come 
into  rapid  favor  because  it  has  not  increased  in  cost  so  much 


208  COAL  MINING  COSTS 

as  the  timber  and  because  it  gives  a  more  permanent  job  and 
is  fireproof  and  more  or  less  watertight. 

The  average  timber  lining  lasts  from  12  to  15  yr.  and  in 
6  to  8  yr.  it  becomes  necessary  to  replace  individual  timbers 
and  sections  of  lining,  causing  temporary  shutdowns.  In  the 
life  of  a  mine  of  any  considerable  size,  say  30  yr.,  it  will  be 
necessary  to  re-timber  the  whole  shaft  at  least  once  besides 
making  many  minor  repairs.  The  cost  of  re-timbering  a  shaft 
(owing  to  the  removal  of  the  old  lining,  and  the  increasing 
price  of  timber  and  labor)  will  be  much  higher  than  the  cost 
of  the  original  lining;  to  this  direct  cost  must  be  added  the 
loss  of  income  from  the  mine  during  the  time  of  repairs. 

The  chief  advantages  of  timber  lining  are  its  lower  first 
cost,  greater  speed  in  placing  it  and  the  fact  that  timber  is 
better  adapted  to  the  rectangular  or  square  form  of  shaft. 
Timber  lining  can  be  placed  in  about  one-third  the  time  that 
concrete  can  which  at  the  average  rate  of  sinking  would 
amount  to  a  saving  in  time  of  about  13  days  for  each  100  ft. 
of  shaft. 

A  comparison  of  a  hoisting  shaft  and  an  air  shaft  of  the 
usual  design  and  with  timber  lining,  with  corresponding  ellip- 
tical and  circular  concrete-lined  shafts,  will  present  an  example 
which  will  closely  approximate  the  conditions  obtained  in  the 
mining  region  of  western  Pennsylvania;  for  other  localities, 
the  local  conditions  governing  the  cost  can  be  substituted.  The 
figures  are  as  of  1905. 

Two  shafts  of  the  United  States  Coal  &  Coke  Co.  (a  sub- 
sidiary company  of  the  United  States  Steel  Corporation)  at 
Tug  river,  West  Virginia,  were  sunk  through  one  seam  of  coal 
at  100  ft.  in  depth,  and  continued  through  a  second  seam  at 
175  ft.  depth.  The  air  shaft  was  14  ft.  2  in.  on  the  short  axis, 
by  20  ft.  on  the  long  axis.  The  shaft  was  lined  for  a  depth  of 
45  ft.  in  order  to  shut  off  the  surface  water.  The  concrete 
was  12  in.  thick. 

The  main  hoisting  shaft  was  17  ft.  4  in.  on  the  short  axis 
by  33  ft.  on  the  long  axis,  the  concrete  being  12  in.  thick  at 
the  sides  and  18  in.  thick  at  the  ends.  It  was  a  four-compart- 
ment shaft,  including  a  downcast  airway,  two  hoisting-ways 
and  a  pipe-way.  It  was  concreted  throughout  on  account  of 
the  downcast  air-way  and  the  desire  to  shut  off  all  the  water 


SHAFT  SINKING 


209 


COSTS  OP  TWO  SHAFTS  FOR  THE  U.  S.  C.  &  C.  Co.  AT  TUG  RIVER,  1905 

Main  Shaft 


Elliptical 

RECTANGULAR 

Timber-lined 

Concrete-lined 

Concrete, 
Excavation, 
Timber, 

per  foot  of  depth, 
per  foot  of  depth, 
per  foot  of  depth. 

4|    cu.  yd. 
13.5cu.  yd. 
90      ft.  B.M. 

5.9  cu.  yd. 
15      cu.  yd. 
80     ft.  B.M. 

12  cu.  yd. 
500  ft.  B.M. 

Cost  per  Foot  of  Depth 


Concrete,      $9  00  per  cu  yd 

$40  50 

$53  10 

Excavation,    5  .  50  per  cu.  yd. 
Timber,         60.  00  per  M  

74.25 
5.40 

$66.00 
30.00 

82.50 
4.80 

Total  cost  of  main  shaft.  .  . 

$120.15 

$96.00 

$140.40 

Air  Shaft 


Elliptical 

RECTANGULAR 

Timber-lined 

Concrete-lined 

Concrete, 
Excavation, 
Timber, 

per  foot  of  depth, 
per  foot  of  depth, 
per  foot  of  depth. 

3  cu.  yd. 
8  cu.  yd. 
70  ft.  B.M. 

8  cu.  yd. 
450ft.  B.M. 

4.3cu.  yd. 
9.9cu.yd. 
70     ft.  B.M. 

Cost  per  Foot  of  Depth 


Concrete,      $10.00  per  cu.  yd. 

$30  00 

$43  00 

Excavation,      6  .00  per  cu  .  yd  . 
Timber,          61  .00  per  cu.  yd. 

48.00 
4.20 

$48.00 
27.00 

59.40 
4.20 

Total  cost  of  air  shaft  

$82.20 

$75.00 

$106.60 

210  COAL  MINING  COSTS 

in  the  rocks;  this  was  successfully  done.  The  cross  buntons 
were  held  by  cast-iron  boxes  built  into  the  concrete,  but  these 
boxes  were  probably  unnecessary.  In  the  hoisting  shaft  an 
average  progress  of  16  ft.  per  week  was  made,  20  ft.  being  the 
maximum.  The  total  excavation  in  this  case  amounted  to 
21  cu.  yd.  per  ft.  of  depth. 

A  paddle  concrete-mixer  was  placed  at  the  head  of  the 
shaft,  and  the  concrete  was  lowered  directly  from  the  dis- 
charge spout  of  the  mixer  without  further  handling.  While 
one  bucket  was  lowering,  the  other  was  filling;  thus  no  time 
was  lost  in  delivering  concrete  to  the  placing  gang.  All  form 
work  was  done  at  night,  and  the  concreting  on  the  day  shift. 
The  forms  were  built  in  5-ft.  vertical  sections  and  were  used 
repeatedly. 

The  cost  of  labor,  mixing,  placing  forms,  lumber,  carpenters, 
hoisting  engineers,  oil,  waste  supplies,  sundries  and  superin- 
tendence amounted  to  about  $4  to  $5  per  yd.  depending  upon 
the  size  of  the  shaft  and  the  thickness  of  the  concrete.  To 
this  must  be  added  the  cost  of  materials. 

The  table,  given  herewith,  illustrates  the  relative  cost  of 
elliptical  concrete-lined  shafts,  and  also  the  usual  rectangular 
shaft  lined  with  concrete  and  with  timber.  It  illustrates  the 
great  economy  of  the  more  permanent  concrete  shaft. 

An  example  of  an  air  and  main  shaft,  each  200  ft.  in  depth, 
would  represent  an  outlay  of  $34,200  for  a  timber-lined  rect- 
angular shaft;  the  equivalent  elliptical  concrete-lined  shaft 
would  cost  $40,440,  both  figures  including  all  materials.  The 
difference  due  to  the  increased  cost  of  $6240  is  more  than  off- 
set by  the  fact  that  it  would  take  over  $15,000  to  re-timber 
both  shafts,  not  to  mention  the  loss  of  time  and  repairs. 

Rates  of  progress. — The  best  record  in  shaft  sinking  up  to 
1907  was  made  at  the  Dixon-Pocahontas  Mine  in  West  Virginia. 
This  shaft  at  Olmstead,  W.  V.,  was  completed  in  nine  work- 
ing weeks,  although  not  in  nine  weeks*  continuous  work,  as 
an  explosion  in  which  four  men  were  killed  disorganized  the 
forces  and  interrupted  th«  work.  This  work  was  practically 
free  from  water  except  the  last  50  ft.  when  a  No.  7  Cameron 
sinking  pump  was  required,  but  after  the  sump  was  com- 
pleted a  No.  10  Cameron  was  installed  for  temporary  use. 

The  shaft  is  14  X  22  ft.  and  180  ft.  deep  to  coal  and  was 


SHAFT  SINKING  211 

9  weeks  in  sinking.  It  is  timbered  with  8  X  10  in.  wall  plates 
and  6  X  10  in.  buntons  and  lagged  with  2-in.  plank.  The  tim- 
bers have  bearing  sets  about  30  ft.  apart  while  the  other  sets 
are  spaced  on  5-ft.  centers. 

The  following  is  the  equipment  installed:  Two  50-hp.  Erie 
City  Economy  and  one  40-hp.  Atlas  internally  fired  boilers,  a 
six-drill  compressor,  a  10  X  12  in.  Exeter  hoist  having  a  4-ft. 
drum  and  %-in-  plow-steel  cable,  a  70-in.  fan  coupled  to  a 
10-hp.  engine,  and  a  blacksmith  shop  outfit. 

The  work  was  carried  down  with  three  8-hr,  shifts  for 
shaft  men  and  two  12-hr  shifts  for  outside  men. 

The  number  of  men  and  their  wages  was  as  follows:  1 
shift  foreman,  $3;  3  drill  runners,  each  $2.50;  3  helpers,  each, 
$2.25 ;  8  muckers,  each,  $2 ;  1  blacksmith,  $2.75 ;  1  hoister,  $2 ;  1 
compressor  man,  $2;  1  carpenter,  $2.50;  2  helpers,  each  $1.75. 

The  coal  for  this  9  weeks*  work  at  $1.25  per  ton  cost  $460. 
The  cost  of  installing  the  sinking  plant  was  about  $1000  while 
that  of  dynamite  was  $700.  Hence,  exclusive  of  machinery, 
the  risk  from  accidents  and  the  cost  of  moving  the  excavated 
material  after  dumping  the  buckets,  we  may  get  a  close  esti- 
mate of  the  cost  of  this  work  as  follows : 

Installing  plant $1000 

Coal 460 

Dynamite 700 

Labor  per  day,  $85  for  54  days 4590 

Timber  for  shaft  at  $20  per  M,  60  M 1200 

Miscellaneous  expenses 1000 


Total $8950 

This  is  practically  an  expense  of  $9000  for  180  ft.  of 
14  X  22  ft.  shaft,  a  net  cost  of  $50  per  foot. 

The  speed  of  sinking  will  be  governed  by  the  quality  of 
the  rock,  size  and  shape  of  the  shaft,  amount  of  water  present, 
class  of  labor  available  and  the  efficiency  of  the  plant.  The 
fastest  sinking  on  record  up  to  1909  was  on  the  Rand  in 
Transvaal,  South  Africa.  These  shafts  are  sunk  9  X  26  ft. 
in  the  rock  and  the  work  is  carried  in  three,  8-hr,  shifts  per 
day,  seven  days  in  the  week.  The  labor  is  cheap  and  can  be 
used  under  conditions  that  the  white  man  will  not  work  under 
so  the  shafts  are  filled  up  with  every  man  that  is  possible  to 
use.  The  records  made  at  some  of  these  shafts  are  given  in 


212 


COAL  MINING  COSTS 


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SHAFT  SINKING  213 

the  accompanying  table,  in  which  it  will  be  noted  that  the 
average  progress  is  about  135  ft.  per  month  with  a  record  sink- 
ing of  213  ft.  in  one  month. 

Progress  in  this  country  up  to  1909  was  under  normal  con- 
ditions at  the  rate  of  60  to  80  ft.  per  month  for  the  average 
timbered  shaft,  though  the  advent  of  improved  drills  since 
that  time  has  much  increased  this  rate.  Faster  records  are 
made  in  the  Middle  West  where  the  soft  shales  are  easy  to 
drill  and  shoot.  Speed  of  7  ft.  per  day  was  made  at  a  shaft 
near  Atchison,  Kan.,  in  1902  which  was  regarded  as  rapid  sink- 
ing at  that  time  though  it  is  not  known  how  long  this  was 
maintained.  In  1909  a  record  of  138  ft.  was  made  in  one  month 
on  a  17-ft.  circular  shaft  through  rock  that  was  quite  hard 
but  broke  readily  on  the  New  York  Aqueduct  which  was  fasi 
sinking  at  that  time,  if  not  a  record.  The  particulars  of  this 
shaft  will  be  found  in  the  accompanying  table. 

The  accompanying  tables  give  the  dimensions  of  a  number 
of  shafts  and  the  progress  made  in  them.  Wherever  obtainable, 
the  nature  of  the  rock  penetrated  and  the  cost  per  foot  is 
given.  The  figures  were  obtained  from  various  articles  in  the 
technical  papers,  and  from  the  proceedings  of  various  mining 
Institutes.  Some  of  the  South  African  data  were  taken  from 
the  "Deep  Level  Mines  of  the  Rand,"  by  G.  A.  Denny, 
1902. 

Most  of  the  shafts  sunk  on  the  Witwatersrand,  South 
Africa,  have  five  compartments.  Some  are  sunk  on  the  dip  of 
the  ledge  which  varies  from  70  to  35  deg.  in  different  mines, 
the  tendency  being  for  the  steeper  dips  to  flatten  to  about 
35  deg.  as  the  mines  get  deeper.  But  most  of  the  shafts  are 
vertical.  Some  new  vertical  shafts  that  have  been  started  for 
working  the  deep  levels  have  seven  compartments. 

Most  of  the  sinking  is  in  quartzite  and  Leslie  Simson,  a 
•graduate  of  the  University  of  California,  held  the  world's 
record  in  1907,  for  sinking  a  vertical  shaft  7  X2$  ft.  in  quart- 
zite at  the  rate  of  203  ft.  in  one  month.  The  average  rate  of 
sinking  eight  shafts  on  the  Consolidated  Gold  Fields  mines, 
7  X  28  ft.,  to  depths  varying  from  1300  ft.  to  3700  ft.  was  kept 
at  about  100  ft.  per  month  and  the  average  cost  $134  per  foot. 
Two  of  these  shafts,  over  3000  ft.  deep,  were  each  sunk  at 
the  rate  of  152  ft.  per  month  for  six  consecutive  months. 


214  COAL  MINING  COSTS 

The  time  on  the  work  was  subdivided  as  follows: 

Hours 

. Hand  drilling  with  Kaffirs 3. 07 

Winding  rock 3 . 26 

Winding  men  and  tools 1 . 17 

Blasting 0. 50 

8.00 

The  number  of  hand-drilled  holes  per  shift  is  20  to  21 ; 
dynamite  used  per  foot,  16  to  18  Ib. ;  coal  burned  per  shift,  2400 
to  2800  Ib.;  buckets  of  rock  hoisted,  34  to  36  (about  1  ton 
capacity). 

Cost  figures  cover  a  wider  range  than  progress  figures  and 
are  harder  to  get.  The  cheapest  shaft  on  record  is  the  one 
near  Atchison  referred  to  above,  the  cost  of  which,  as  stated, 
was  $7  per  foot.  This  cost  stands  alone  in  its  glory  as  the 
tabulated  figures  show.  Mr.  Henry  Eawie  published  in  Mines 
and  Minerals  an  itemized  statement  of  the  costs  of  a  shaft 
sunk  in  West  Virginia,  in  1906.  These  ran  as  follows : 

HOIST  SHAFT,  14X22  FT.,  180  FT.  DEEP 

Per  Foot 

Labor,  sinking,  and  timbering $24 . 70 

Plant 5.55 

Superintendence • 

Explosives 3 . 88 

Coal 2.55 

Timber 6. 67 

Miscellaneous ' 5 . 55 


$48.90 

The  sinking  costs  of  a  pair  of  shafts  sunk  in  Western  Penn- 
sylvania a  year  later  were  as  follows: 

HOIST  SHAFT,  13X26  FT.,  422  FT.  DEEP 

Per  Foot 

Labor,  sinking $51 . 00 

Plant 2.40 

Superintendence 4 . 35 

Explosives 2 . 75 

Coal 5.50 

Oil 0.60 

Freight 0. 50 

Miscellaneous 7 . 90 

Total..  .  $75.00 


SHAFT  SINKING  215 

AIR-SHAFT,  13X22  FT.,  383  FT.  DEEP 

Per  Foot 

Labor,  sinking $57 . 50 

Plant 2.40 

Superintendence 4 . 90 

Explosives 3 . 00 

Coal 6.05 

Oil 0.60 

Freight 0.50 

Miscellaneous . .                                           7 . 14 


Total $82.09 

Water  per  minute :    Hoist  shaft,  50  gal. ;  air-shaft,  120  gal. 

Costs  have  risen  greatly  in  the  last  decade  since  no  sub- 
stantial improvements  in  methods  or  machinery  have  been 
made  to  offset  the  increase  in  wages.  Contract  prices  are  not 
generally  obtainable,  as  most  shafts  are  put  down  by  private 
corporations,  but  prices,  high  enough  to  include  a  good  profit 
to  the  contractor  8  or  10  yr.  ago,  would  not  cover  his  costs 
to-day. 

Twenty-five  shafts  ranging  in  depth  from  350  to  over 
1000  ft.  for  a  portion  of  the  New  York  Aqueduct  were  con- 
tracted for  at  prices  ranging  from  $175  to  $350  per  foot. 

Reports  and  contract  forms. — Proper  supervision  of  costs  of 
supplies,  labor  and  progress  of  work  can  only  be  obtained  by 
keeping  a  detailed  daily  report  of  the  sinking  operations  and 
the  accompanying  form,  Figs.  5  and  6,  show  a  very  good 
method  of  doing  this.  These  forms  were  used  by  the  Cotton- 
wood  Coal  Co.,  a  subsidiary  of  the  Great  Northern  R.  E.  on 
a  31  X  9-ft.  five-compartment  shaft  at  Lehigh,  Mont.  The 
forms  are  made  out  in  triplicate  and  copies  forwarded  to  the 
president  and  general  manager  and  one  kept  on  file  at  the 
plant.  The  report  is  self  explanatory  and  can  with  modifica- 
tions be  adapted  to  almost  any  condition. 


216 


COAL  MINING  C08TS 


£& 


TOTAL  FOR  MONT 


FIG.  5. — Front  of  daily  report  form  for  shaft  sinking. 


DAILY  LABOR  STATEMENT 


«TT         101. 


LABOR  CLASSIFICATION 


J&£ 


SURFACt  IASORERS 


FIG.  6. — Back  of  shaft  sinking  report  form. 


SHAFT  SINKING  217 

The  following  is  a  contract  form  used  by  one  shaft  company, 
revised  to  1921 : 

SHAFT  SINKING  CO. 

PITTSBURGH,  PA ,  19. . 


We  hereby  propose  to 

in  accordance  with  the  following  specifications: 

I.  SERVICES  TO  BE  RENDERED 
With  respect,  to  this  work,  we  propose  to 


Upon  your  acceptance  of  this  proposal,  we  will  assemble  at  the  mine  site  as  promptly  as 
possible  a  crew  of  experienced  and  efficient  workmen,  together  with  a  mining  captain,  and 
such  foremen  and  clerical  help  as  are  necessary.  We  will  thereupon  proceed  with  the  pnose- 
cution  of  the  work,  carrying  on  the  same  thereafter  with  all  reasonable  diligence  and  in  a 
skillful  and  workmanlike  manner  until  its  completion,  in  accordance  with  the  plans  and 
specifications  adopted  for  the  work,  subject,  however,  to  any  delays  which  may  be  caused 
by  weather  conditions,  labor  strikes,  accidents,  and  other  causes  beyond  our  control,  or 
changes  in  the  approved  plans  which  may  be  required  by  unexpected  conditions  or  by  your 
instructions.  In  the  prosecution  of  the  work  we  will  be  guided  in  all  respects  by  such 
specific  instructions  as  you  may  give  us  from  time  to  time. 

II.  EQUIPMENT  AND  MATERIALS 

Adequate  machinery,  equipment,  materials  and  supplies  for  the  prosecution  of  the  work 
are  to  be  furnished  and  paid  for  by  you;  but  at  your  request,  or  in  case  of  your  failure 
promptly  to  provide  the  same,  and  in  order  to  prevent  delay  in  the  completion  of  the  work, 
we  will  procure  needful  materials  and  equipment  and  charge  the  cost  thereof  to  you,  the 
amount  of  such  cost  to  be  paid  to  us  by  you  at  the  next  monthly  settlement,  as  hereinafter 
provided. 


III.  COMPENSATION 

As  compensation  for  our  services  performed  in  connection  with  the  said  work,  you  are 

to  pay  us  the  sum  of 

any  such  compensation  (as  well  as  any  expenditures  made  by  us  on  account  of  machinery, 
equipment  or  materials,  as  provided  in  the  preceding  paragraph)  to  be  paid  to  us  as  specified 
in  Paragraph  VI. 


IV.  COST  OF  WORK 

It  is  understood  that  you  are  to  pay  all  the  costs  of  the  work,  including: 
(a)  The  amount  paid  for  all  materials,  machinery,  tools,  and  equipment  furnished  by 
us,  whether  the  same  or  any  part  thereof  be  purchased  for  the  work  or  rented  for  use  therein, 
the  maintenance  and  insurance  thereof,  and  the  cost  of  replacement  of  any  machinery, 
tools  or  equipment  that  may  be  worn  out  or  destroyed;  all  such  machinery  tools  and 
equipment,  except  that  which  may  have  been  rented,  to  belong  to  you  at  the  completion 
of  the  work. 


218  COAL  MINING  COSTS 

(b)  All  amounts  paid  for  labor  and  bonuses  to  workmen,  provided  that  the  scale  of  wages 
and  amount  of  bonuses  shall  be  approved  by  your  engineers. 

(c)  The  cost  of  all  traveling  and  transportation  expenses  of  men  and  of  machinery, 
tools,  equipment  and  materials  supplied  by  us,  including  the  cost  of  return  transportation 
to 

(d)  The  cost  of  liability  and  other  forms  of  insurance  carried  by  us,  and  any  expense 
incurred  in  connection  with  any  accidents  or  damage  to  person  or  property. 

(e)  The  cost  at  salary  rate  of  our  Mining  Superintendent  for  the  time  actually  spent  by 
him  on  work  under  this  contract,  either  in  the  field  or  at  our  office,  and  his  expenses;    and 
any  other  expenditures,  not  herein  specified,  made  by  us  in  the  proper  carrying  out  of  the 
work,  but  not  including  any  overhead  expenses  of  our  home  office. 


V.  INSURANCE 

All  employers'  liability  or  workmen's  compensation  insurance  procured  by  us  for  our 
protection  in  carrying  on  the  work  under  this  contract  shall  be  paid  for  by  you,  unless  you 
shall  elect  to  procure  such  insurance  or  to  carry  such  risks  yourselves,  and  shall  enter  into  a 
satisfactory  contract  with  us,  assuming  and  agreeing  to  pay  all  damages  and  costs  accruing 
through  any  accident  or  injury  to  workmen  or  others  during  the  prosecution  of  the  work, 
and  to  indemnify  and  hold  us  harmless  from  any  liability  or  costs,  including  attorney's 
fees,  on  account  of  the  same. 

VI.  PAYMENTS 

As  promptly  as  possible  after  the  end  of  each  month,  we  will  render  to  you  duplicate 
payrolls,  with  statements  of  all  expenditures  made  by  us  during  such  month,  the  amounts 
shown  by  such  payrolls  and  statements  to  be  paid  to  us  by  you  in  New  York  exchange  on 
or  before  the  fifteenth  day  of  the  month  in  which  the  same  are  rendered,  together  with  our 
compensation. 


VII.  REPORTS 

We  will  render  reports  to  you  monthly  showing  the  progress  of  the  work,  with  any 
rcornmendations  for  changes  in  specifications  adopted  which  it  may  seem  advisable  to  make. 


VIII.  TERMINATION  OF  CONTRACT 

If,  at  any  time,  you  shall  become  dissatisfied  with  the  manner  in  which  the  work  is  being 
conducted  by  us,  or  shall  wish,  for  any  reason,  to  discontinue  the  work,  you  will  be  at  liberty 
to  terminate  our  employment  hereunder  upon  giving  us  ten  (10)  days'  notice  in  wr-ting  of 
your  intention  so  to  do,  and  at  the  expiration  of  said  period  of  ten  (10)  days,  we  will  surrender 
to  you  possession  of  the  work;  provided  that,  in  such  case,  we  shall  be  entitled  to  and  you 
shall  pay  us,  as  our  compensation  for  our  services  hereunder, 


ACCEPTANCE  AND  APPROVAL 

On  acceptance  of  this  proposal  by  you,  and  its  approval  by  the  President,  Vice-President, 
or  Manager  of  the  Mining  Department  of  this  Corporation  (unless  the  proposal  shall  be 
signed  by  one  of  them),  this  instrument  shall  constitute  a  binding  agreement  between  us. 

SHAFT  SINKING  Co. 


By 

Accepte    ,  19 ... 

, Approved ,      19 . 

By..  


SECTION  III 
HAULAGE  COSTS 

With  any  method  of  handling  coal  underground  the  item 
of  haulage  is  a  rather  small  percentage  of  the  total  cost  of 
production  of  a  ton  of  coal.  It  represents,  however,  one  of 
the  items  that  can  be  varied  by  applying  different  methods, 
and  a  material  reduction  in  costs  can  quite  often  be  effected 
by  a  careful  study  of  conditions  and  a  proper  application  of 
equipment  designed  to  do  certain  work. 

The  haulage  system  also  has  an  important  bearing  on  other 
costs  that  occur  in  producing  coal  and  which  are  not  directly 
a  part  of  the  haulage  system.  Because  of  the  high  speed  and 
flexibility  of  its  operation,  it  is  at  times  possible  to  do  intensive 
mining,  thereby  reducing  the  active  area  in  the  mine  with  a 
less  ventilation  cost,  smaller  trackage,  easier  supervision,  and 
in  many  cases  it  permits  of  greater  recovery. 

The  actual  efficiency  of  a  haulage  system,  when  input  power 
is  compared  with  actual  work  in  delivering  coal  from  the  face 
of  the  workings  to  the  shaft  bottom  or  tipple  is  extremely  low, 
considering  the  difficulties  to  contend  with.  While  it  is  well 
to  consider  the  economical  use  of  power  at  all  times,  it  repre- 
sents only  a  small  percentage  of  the  haulage  costs  and  is  not 
generally  susceptible  of  material  improvement. 

The  real  reduction  in  haulage  costs  comes  from  so  arrang- 
ing the  work  that  each  piece  of  equipment  is  operated  to  its 
maximum  capacity  at  all  times.  This  is  the  really  difficult 
problem  to  solve,  as  it  is  interlocked  with  methods  of  mining, 
drainage  and  ventilation,  and  the  solution  can  only  be  worked 
out  by  men  having  an  intimate  acquaintance  with  the  condi- 
tions surrounding  any  specific  problem  under  consideration. 

In  considering  the  item  of  haulage  the  problem  of  obtain- 
ing an  economic  grade  for  each  section  of  the  development  is 
primary  in  importance.  There  are  localities,  however,  where 

219 


220 


COAL  MINING  COSTS 


the  seam  lies  approximately  level  and  entries  or  headings  may 
be  driven  in  any  direction  convenient  for  other  reasons.  The 
engineer  must  then  plan  the  system  of  working  to  give  the 
shortest  haul  with  the  minimum  expense  for  construction  and 
maintenance  of  roadways  and  also  the  least  cost  for  entry 
driving. 

In  localities  where  union  labor  is  employed  a  charge  for 
yardage  is  made  for  driving  the  entries  narrow.    Narrow  work 


FIG.  1. — Plan  for  driving  new  entries  to  reduce  length  of  haulage. 

usually  includes  all  work  less  than  18  ft.  wide.  The  price  per 
linear  yard  of  entry  varies  somewhat  in  different  fields,  as  it 
is  fixed  in  the  district  contracts  made  between  the  operator 
and  the  labor  organizations. 

By  increasing  the  amount  of  entry  driven  to  a  certain 
amount  the  shortest  haul  may  be  obtained.  To  effect  this 
advantage  the  expense  of  constructing  and  maintaining  the 
additional  roadway  and  a  reduction  of  efficiency  through  loss 
of  ton  mileage  per  foot  of  motor  road  operated  must  be  con- 
sidered. 


HAULAGE  COSTS  221 

The  illustration,  Fig.  1,  shows  a  modified  panel  system  of 
mining  which  is  rapidly  being  adopted  in  many  sections.  From 
the  main  heading  A  secondary  headings  B  and  D  are  driven, 
and  from  the  secondary  headings  the  room  or  butt  Headings 
are  driven  on  both  sides.  Partings  are  located  on  the  butt 
headings  just  off  the  secondary  heading. 

In  gathering  from  rooms  on  butt  heading  N  the  coal  is 
hauled  away  from  its  ultimate  destination;  in  gathering  from 
rooms  on  heading  M  the  coal  is  hauled  toward  its  ultimate  des- 
tination. In  order  to  eliminate  this  back  haul  the  secondary 
heading  X,  indicated  in  the  sketch  by  dotted  lines,  could  be 
driven  and  all  coal  from  the  territory  of  N  and  P  taken  out  over 
a  roadway  located  through  this  heading.  By  driving  this  addi- 
tional heading  the  mine  may  develop  sooner  in  that  particular 
section,  but  the  usual  rush  for  coal  at  an  immediate  low  cost 
too  often  induces  a  method  of  working  which  results  in  a  sacrifice 
of  future  profits. 

The  cost  for  yardage  in  driving  headlong  X  combined  with 
many  other  items  will  be  an  additional  expense  to  be  prorated 
over  the  cost  of  producing  the  tonnage  from  the  territory  served. 
The  items  of  expense  for  heading  X  alone  may  be  expressed 
by  formula?. 

Let  y  =  length  of  heading  X  in  feet; 

c  =  yardage  cost  of  driving!  yd.  of  entry  expressed  in 

dollars ; 

n  =  number  of  entries  on  the  heading; 
Ci  =  total  cost  of  yardage  for  X; 

C2  =  cost  of  construction  of  overcasts,  brattices,  doors,  etc.; 
Ca  =  cost  of  construction  of  main  haulage  track,  exclusive 
of  switches,  minus  the  salvage  value  of  the  material 
when  released; 

£4  =  cost  of  maintaining  the  roadway,   stoppings,  doors, 
overcasts,    etc.,    during   the   period   coal   is   being 
hauled; 
E  =  the  value  of  the  loss  of  efficiency  from  a  reduced  ton 

mileage  per  foot  of  main  motor  road  operated. 
Then: 

Ci"7;    ••,;•. (1) 

and 

(2) 


222  COAL  MINING  COSTS 

where  C  =  total  expense  resulting  from  having  driven  the  heading 
X. 

The  value  of  E  is  really  a  function  of  C  although  difficult  of 
calculation.  It  is  easy  to  see,  however,  that  if  heading  X  is 
omitted  the  traction  on  roadway  B  will  be  doubled  and  the 
combined  tonnage  (T)  from  butt  headings  N  and  P  will  have  an 
average  back  haul  of  Z/2  ft. 

Let  V  =  cost  of  hauling  1  ton  1  mile  underground,  expressed  in 

dollars; 

K  —  total  cost  of  back  haul  in  dollars. 
Then: 

TVZ        TVZ 

•*•    *   *J  L    \  £J  , 


5280X2     10560* 

The  advantages  of  driving  the  additional  heading  are  now 
expressed  by  the  value  of  K  and  the  disadvantage  expressed 
by  the  value  of  C.  If  other  factors  are  not  considered,  the 
ratio  of  these  values  determines  the  ultimate  plan  to  be  pursued. 
The  values  for  the  assumed  variables  must  be  chosen  only  after 
accumulating  and  digesting  all  pertinent  information  obtainable 
in  the  same  operating  field  or  in  a  field  supposedly  similar.  The 
value  of  the  results  will  depend  upon  the  competent  judgment 
employed  in  the  calculation,  and  the  equations  simply  represent 
a  plan  of  reasoning  and  investigation  that  should  be  under- 
taken before  adopting  a  decisive  system  of  operation. 

Tractive  effort,  drawbar-pull  and  rating  of  mine  motors.  — 
A  clear  grasp  of  the  relative  economy  and  comparative  cost 
of  operation  of  the  various  types  of  under-ground  haulage 
motors,  necessitates  a  thorough  understanding  of  the  theories 
governing  their  operation.  There  are  several  distinctly  separate 
factors  concerned  in  locomotive  haulage,  which  may  briefly  be 
described  as  follows: 

The  tractive  effort  of  a  locomotive  is  the  force  exerted  by 
the  motor  at  the  circumference  of  the  driving  wheels.  In  a 
well-designed  machine,  the  power  of  the  motor  is  such  as  to 
equal  the  greatest  adhesion  of  the  wheels  to  the  rails,  which 
adhesion  must  of  necessity  limit  the  possible  tractive  effort  of 
the  machine.  This  tractive  effort,  as  thus  limited  by  the 
adhesion  of  the  wheels  to  the  rails,  is  therefore  the  force  avail- 
able to  move  the  entire  load,  including  the  locomotive  and  the 
trip  it  hauls. 


HAULAGE  COSTS  223 

The  drawbar  pull  is  that  portion  of  the  tractive  effort  that 
is  employed  to  move  the  trip  attached  to  the  locomotive.  Thus 
the  hauling  capacity  of  a  locomotive,  as  represented  by  the 
possible  drawbar  pull,  is  always  less  than  the  maximum  tractive 
effort  the  locomotive  can  exert,  by  an  amount  equal  to  the 
force  required  to  move  the  locomotive  itself. 

It  should  also  be  clear  that  the  drawbar  pull  is  always 
equal  to  the  total  resistance  offered  by  the  trip  hauled,  whether 
the  locomotive  is  taxed  to  its  full  capacity  or  not.  In  other 
words,  the  drawbar  pull  in  any  event  is  limited  by  the  resist- 
ance of  the  trip  hauled. 

The  track  resistance  is  the  frictional  resistance  offered  by 
the  entire  moving  load  and  is  estimated  in  pounds  per  ton 
of  load. 

The  grade  resistance  is  the  gravity  pull  of  a  load  resting 
on  an  inclined  track  and  is  equal  to  the  weight  of  the  load 
multiplied  by  the  percentage  of  grade  expressed  decimally. 
Hence,  grade  resistance  is  always  20  Ib.  per  ton  for  each  per 
cent  of  grade,  since  2000  Ib.  equals  one  ton  and  0.01  X  2000 
=  20  Ib. 

The  actual  drawbar  pull,  in  any  case,  is  equal  to  the  sum 
of  the  track  resistance  and  grade  resistance  of  the  load  hauled. 
For  purposes  of  estimate,  it  is  common  practice  to  assume  the 
possible  drawbar  pull  as  varying  from  20  to  30  per  cent  of 
the  weight  of  the  locomotive  resting  on  the  drivers,  according 
to  the  condition  of  the  rails  and  the  kind  of  wheels  used.  But, 
the  actual  drawbar  pull  must  be  equal  to  the  total  resistance 
(track  and  grade  resistances)  of  the  cars,  which  is  estimated  in 
pounds  per  ton  of  load  hauled. 

Track  resistance,  in  mining  practice,  will  vary  from  20  to 
40  Ib.  per  ton,  while  grade  resistance  is  20  Ib.  per  ton  for  each 
per  cent  of  grade.  Assuming  a  level  road  and  a  track  resistance 
of,  say,  25  Ib.  per  ton  of  load,  the  load  that  a  6-ton  locomotive 
will  haul  on  a  level  track,  the  drawbar  pull  being  2400  Ib.,  is 
2400  -r-  25  =  96  tons. 

For  the  sake  of  illustration,  let  is  be  required  to  find  the 
maximum  load  a  6-ton  mine  locomotive  will  haul  up  a  2y2- 
per  cent  grade,  assuming  a  track  resistance  of  30  Ib.  per  ton. 
In  this  case,  the  grade  resistance  is  2%  X  20  ==  50  Ib.  per  ton. 
The  track  resistance  being  30  Ib.  per  ton  makes  the  total 


224  COAL  MINING  COSTS 

resistance  of  the  load  hauled  50  -|-  30  =  80  Ib.  per  ton.  Then, 
taking  the  drawbar  pull  as  approximately  one-fifth  of  the 
weight  of  the  locomotive,  or  V5  (6  X  2000)  =  2400  Ib.,  the  maxi- 
mum load  this  locomotive  will  haul  on  such  a  grade  is  2400 
-l-  80  =  30  tons. 

Generally,  tractive  effort  is  preferable  to  speed  in  a  mine 
motor  of  any  discription.  As  an  example,  a  15  hp.  electric 
motor  rated  at  9  miles  per  hour  requires  50  per  cent  more 
power  to  start  a  certain  load  than  a  10  hp.  motor  rated  at 
6  miles  per  hour. 

Probably  the  best  speed  to  obtain  the  maximum  economy 
in  operation  is  about  6  miles  per  hour  for  motors  up  to  10  tons 
capacity  and  7  to  8  miles  per  hour  for  those  of  larger  capacity. 
The  actual  running  speed  of  nearly  all  electric  motors  is  from 
30  to  40  per  cent  higher  than  the  rated  speed  after  the  load  is 
started. 

Mine  motors  are  rated  at  the  end  of  the  armature  shaft. 
The  following  is  a  short  and  convenient  formula  for  determin- 
ing the  horsepower  output  at  drawbar  of  any  motor : 

„                     Tractive  effort  in  Ib.X speed  in  mi.  per  hr. 
Horsepower  = ^ 

Oi  O 

The  tractive  effort  equals  one-sixth  the  weight  of  the  motor 
and  the  effective  wattage  per-mile-hour  is  double  the  tractive 
effort.  The  horsepower  required  to  operate  a  7%  ton  motor 
at  6  miles  per  hour  when  running  at  90  per  cent  efficiency  is : 

6(2X2500  tractive  effort)     QQ 

n  nn   a*  • -=33,333  watts  •*•  746  =45  h.p. 

0.90  efficiency 

The  accompanying  table  gives  the  drawbar  pull  and  haulage 
capacity  of  electric  mine  motors  of  from  3  to  30  tons  weight 
working  on  grades  up  to  10  per  cent,  as  compiled  by  the 
Jeffrey  Manufacturing  Co. 

Where  possible  an  actual  test  should  be  made  in  order  to 
determine  the  average  frictional  resistance  of  the  cars.  This 
test  can  be  made  by  pulling  a  car  at  a  constant  speed  on  a 
level  track  or  on  a  track  of  a  known  grade  and  measuring 
the  pull  by  means  of  a  spring  balance.  Another  method  which 
will  give  the  approximate  friction  is  to  find  a  grade  down 
which  a  car  will  coast  slowly  at  a  constant  speed  without  tend- 


HAULAGE  COSTS 


225 


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HAULAGE  COSTS  227 

ing  to  increase  or  decrease  its  velocity.  If  this  grade  is  say 
1  per  cent  then  the  fractional  resistance  of  the  car  would  be  1 
per  cent  of  2000  or  20  Ib.  per  ton. 

When  the  frictional  resistance  of  the  cars  is  not  given  it 
should  be  assumed  at  30  Ib.  per  ton  unless  they  are  newly 
equipped  with  roller  bearings  of  an  approved  make.  Roller 
bearings  when  installed  and  looked  after  properly,  will  no 
doubt  give  frictional  resistance  ranging  from  15  to  20  Ib.  per 
ton.  However,  a  number  of  roller  bearings  which  had  been 
neglected  and  not  properly  lubricated  were  once  tested  and 
showed  the  average  resistance  of  several  cars  was  36  Ibs. 
per  ton. 

On  account  of  the  short  wheel  base  of  a  mine  car  the  fric- 
tion may  be  considerably  higher  when  pushed  than  when 
pulled.  When  a  string  of  cars  are  pushed  they  are  liable  to 
wobble  more  or  less,  and  considerable  binding  of  the  flanges 
against  the  rails  may  take  place.  When  the  cars  are  pulled 
they  are  stretched  out  straight  and  but  little  wobbling  will 
take  place.  The  frictional  resistance*  of  new  cars  will  largely 
depend  upon  the  type  of  bearing,  while  for  the  old  cars  it 
will  be  decidedly  influenced  by  the  manner  in  which  the  bear- 
ings are  kept  up. 

The  locomotive  resistances  will  range  from  12  to  20  Ib.  per 
ton.  It  is  safe  to  take  15  Ib.  as  an  average  since  the  friction 
of  the  locomotive  is  such  a  small  percentage  of  the  total  trac- 
tive effort,  and  a  change  of  several  pounds  in  either  direction 
will  not  affect  the  weight  of  the  locomotive  appreciably,  and 
the  effect  on  the  capacity  of  the  motors  will  be  negligible. 

The  effect  of  the  frictional  resistance  of  the  load  will  vary 
with  the  length  and  severity  of  the  grades.  If  the  track  is 
practically  level  throughout,  then  a  small  change  in  the  fric- 
tional resistance  may  have  considerable  effect  on  both  weight 
and  equipment.  If,  however,  the  grades  are  long  and  severe 
the  effect  will  be  small. 

When  a  locomotive  is  operating  at  a  constant  speed  on  a 
straight,  level  track,  the  drawbar  pull  available  for  hauling  a 
trailing  load  (provided  there  is  sufficient  motive  power)  is 
limited  only  by  the  adhesion  that  can  be  obtained  between 
the  driving  wheels  and  the  rails.  When  starting,  the  drawbar 
pull  available  is  reduced,  depending  upon  the  rate  of  accelera- 


228  COAL  MINING  COSTS 

tion.  As  this  rate  is  seldom  more  than  0.2  to  0.25  mi.  per  hour 
per  second,  the  drawbar  pull  will  be  reduced  from  19  to  24  Ib. 
for  each  ton  weight  of  locomotive. 

If  there  are  no  grades  the  weight  of  the  locomotive  will 
be  affected  considerably  by  the  rate  of  acceleration.  With 
heavy  grades,  however,  the  acceleration  will  have  little  effect 
since  the  rate  can  be  kept  low  if  it  becomes  necessary  to  start 
on  the  heavy  grade.  Accordingly  with  the  low  rate  of  accelera- 
tion common  to  mine  service  this  factor  can  be  considered 
negligible  as  regards  the  weight  of  the  locomotive,  in  view  of 
the  fact  that  a  greater  percentage  of  adhesion  can  be  allowed 
for  starting  by  the  use  of  sand. 

It  has  been  found  in  practice  that  with  cast-iron  wheels  a 
running  drawbar  pull  equivalent  to  an  adhesion  of  20  per  cent 
of  the  weight  on  the  drivers  can  be  obtained  with  clean  dry 
rails  on  level  track,  without  the  use  of  sand.  A  steel-tired  or 
rolled-steel  wheel  seems  to  obtain  a  better  grip  on  the  rails, 
and  a  drawbar  pull  equivalent  to  an  adhesion  of  25  per  cent 
can  be  obtained  under  the  same  conditions.  When  starting 
heavy  trips  and  when  on  steep  grades  it  is  permissible  to  use 
sand,  in  which  case  a  drawbar  pull  equivalent  to  25  to  30 
per  cent  for  cast-iron  wheels  and  30  to  33V3  per  cent  for  steel 
wheels  can  be  expected. 

Where  grades  are  short  the  higher  rates  of  adhesion  may 
be  used,  but  for  long  grades  it  is  not  the  best  practice. 
Dynamometer  tests  have  given  adhesion  values  as  high  as  40 
to  45  per  cent  by  the  use  of  sand.  The  average  of  the  tests 
was,  however,  much  lower  so  that  it  is  not  good  practice  to 
count  on  such  high  values.  These  high  percentages  require 
the  liberal  use  of  sand  on  both  rails,  a  practice  which  should 
not  be  encouraged  as  the  sand  increases  the  fractional  resist- 
ance of  the  locomotive  and  cars,  and  may  work  into  the  bear^ 
ings  and  gears. 

Where  no  grades  exist  the  weight  of  the  locomotive  should, 
therefore,  be  five  times  the  drawbar  pull  for  cast-iron  wheels 
and  four  times  for  steel  wheels,  unless  the  rate  of  acceleration 
is  such  that  additional  weight  is  required.  When,  however, 
a  locomotive  with  a  trailing  load  is  ascending  a  grade  the  draw- 
bar pull  is  necessarily  greater  than  that  required  to  overcome 
the  friction  of  the  trailing  load  as  the  weight  of  the  load  has 


HAULAGE  COSTS  229 

to  be  lifted  up  the  grade.  For  every  1  per  cent  grade  20  Ib. 
per  ton  should  be  added  to  the  drawbar  pull  required  on 
straight  level  track  since  20  is  1  per  cent  of  2000  Ib. 

The  effect  of  grade  on  the  locomotive  as  well  as  on  the  load 
must  be  considered.  The  heavier  the  grade  the  less  will  be 
the  drawbar  pull  of  the  motor.  This  becomes  evident  when 
an  abnormal  grade  is  considered  on  which  a  motor  will  be 
barely  able  to  .propel  itself,  and  if  any  trailing  load  is  added 
the  wheels  will  slip.  The  greater  tendency  for  the  wheels  to 
slip  on  a  grade  is  due  to  the  increased  tractive  effort  necessary 
to  propel  the  motor  itself  and  the  weight  transfer  due  to 
grade. 

The  weight  transfer  due  to  grade  will  depend  on  the  wheel 
base  and  height  of  the  center  of  gravity.  With  a  short  wheel 
base  and  a  high  center  of  gravity  the  weight  transfer  will  be 
considerable.  The  modern  mine  motor,  however,  is  constructed 
with  a  low  center  of  gravity  and  a  fairly  long  wheel  base  so 
that  with  the  ordinary  grades  encountered  the  weight  transfer 
is  not  serious. 

The  weight  transfer  due  to  height  of  drawbar  will  also  effect 
the  drawbar  pull  if  the  wheel  base  is  short  and  the  drawbar 
high.  In  Fig.  2  a  represents  the  height  of  drawbar  and  6  the 


FIG.  2. — Influence  of  drawbar  height  on  weight  transfer  of  car. 

wheel  base.  A  horizontal  force  at  the  drawbar  will  act  as  a 
bell  crank,  one  of  whose  arms  is  a  and  the  other  &.  If  the 
horizontal  force  represented  by  the  drawbar  pull  is  in  the  direc- 
tion of  the  arrow,  an  upward  force  will  be  exerted  at  x  and  a 
downward  force  at  y.  If  the  drawbar  pull  is  represented  by  D 
then  the  moment  about  y  will  be  Da.  This  moment  divided  by  b 
will  give  the  lifting  force  at  x.  The  adhesion  of  the  wheel  x 
will  of  course  be  lessened  by  the  lifting  force. 

Assume  a  locomotive  weighing  12  tons  with  a  wheel  base  of 


230  COAL  MINING  COSTS 

5  ft.  and  the  height  of  drawbar  10  in.  At  25  per  cent  adhesion 
Da  will  be  6000  X  if  =  5000  ft.  Ib.  The  upward  pull  at  x  will 
be  5000  ~  5  =  1000  Ib.  The  normal  weight  at  x  is  12,000  Ib. 
therefore  the  weight  when  the  drawbar  pull  is  6000  Ib.  will  be 
11,000  Ib.  so  that  the  adhesion  will  be  3000/110oo  =  27.2  per  cent. 
It  is  thus  to  be  seen  that  with  the  ordinary  height  of  drawbar 
and  wheel  base  the  drawbar  pull  is  not  seriously  affected.  In 
metal  mining  a  much  higher  drawbar  is  sometimes  required  so 
that  the  wheel  base  must  be  lengthened  to  lessen  the  tilting 
effect. 

By  tying  the  axles  together  by  means  of  side  rods,  chains 
or  gears  the  effect  of  weight  transfer  due  to  grade  and  height 
of  drawbar  can  be  eliminated,  but  these  devices  have  not 
proven  successful  from  an  operating  standpoint. 

Ordinary  mine  cars  will  require  from  30  to  40  Ib.  pull  per 
ton  to  move  them  on  a  level  track.  The  horizontal  tractive 
resistance  of  modern  cars  in  good  condition  may  drop  down  as 
low  as  20  Ib.  per  ton,  but  it  is  not  safe  to  figure  much  less  than 
30  to  35  Ib.  per  ton  on  the  level.  For  each  1  per  cent  grade 
against  the  load,  20  Ib.  per  ton  must  be  added.  For  example, 
if  a  car  will  move  on  30  Ib.  per  ton  pull  on  a  level,  it  will 
require  30  plus  20  or  50  Ib.  pull  against  a  1  per  cent  grade, 
or  30  plus  20  plus  20  or  70  Ib.  pull  against  a  2  per  cent  grade. 

A.  M.  Wellington  found  that  it  required  5  Ib.  per  ton  to 
pull  a  loaded  freight  car  and  7  Ib.  per  ton  to  pull  an  empty 
freight  car  over  a  level  railroad  track  at  10  mi.  per  hr. 

R.  Van  A.  Norris  made  980  tests  some  years  ago,  which 
gave  a  resistance  of  26  Ib.  per  ton  for  20  loaded  mine  cars 
and  42  Ib.  per  ton  for  20  empty  mine  cars  of  the  same  size 
over  the  same  track,  traveling  4%  mi.  per  hr.  One  reason  for 
the  wide  difference  between  the  two  sets  of  experiments  men- 
tioned is  found  in  the  size  of  the  car  wheels. 

The  freight  cars  had  32-in.  diameter  wheels,  while  the  mine 
cars  had  16-in.  diameter  wheels,  thus  making  it  possible  for 
the  former  to  ride  over  the  inequalities  in  the  roadbed  with 
more  ease  than  the  latter. 

Mr.  Norris 's  experiments  brought  out  another  important 
matter  not  usually  mentioned  in  haulage  articles ;  namely,  that 
it  requires  more  power  per  ton  to  haul  short  trains  of  mine 
cars  than  longer  ones.  This  is  of  added  importance  to  those 


HAULAGE  COSTS  231 

mines  which  have  adopted  the  two-unit  locomotive  in  order 
to  pull  heavier  loads. 

One  factor  that  probably  reduces  the  ton  resistance  in 
longer  trips  is  that  the  forward  cars  clean  the  track;  and 
another  is  that  the  greater  momentum  of  longer  trips  aids  in 
keeping  the  cars  moving. 

Professor  Baker  made  some  experiments  on  clean  and  dirty 
tracks,  the  results  of  which  briefly  were  as  follows:  On  a 
perfectly  clean  track  the  resistance  was  19  Ib.  per  ton;  on  the 
same  track  with  %  in.  of  fine  dust  the  resistance  was  28  Ib. 
per  ton;  while  with  %  in.  of  powdered  stone  the  resistance 
was  40  Ib.  per  ton.  This  should  comfort  superintendents  in 
the  soft-coal  mines  who  begrudge  spending  money  on  road 
cleaning,  because  when  they  clean  the  haulage  entries,  they  not 
only  lessen  the  dangers  from  explosions,  but  decrease  the  cost 
of  haulage. 

It  is  also  a  lesson  for  some  anthracite  superintendents,  for 
fine  anthracite  is  nearly  equal  to  sand  in  offering  resistance 
to  traction  effort. 

A  gasoline  motor  will  exert  a  drawbar  pull  equal  to  one- 
fifth  its  weight  in  pounds  when  working  on  a  level  and  on  dry 
rail  of  proper  weight  for  the  motor,  but  from  this  drawbar 
pull  exerted  on  a  level  we  deduct  1  per  cent  of  the  weight  of 
the  motor  in  pounds  for  each  1  per  cent  grade.  For  example, 
a  5-ton  motor  will  exert  a  tractive  effort,  or  drawbar  pull,  on 
the  level  of  one-fifth  of  10,000  Ib.,  or  2000  Ib.  Against  a  1  per 
cent  grade  we  would  have  2000  Ib.  less  1  per  cent  of  10,000  Ib., 
or  2000  less  100  Ib.,  or  1900  Ib.,  net;  or  1800  Ib.,  net  against  a 
2  per  cent  grade. 

It  is  not  only  advisable  for  the  operator  to  prefer  the 
locomotive  with  the  higher  tractive  effort  and  low  speed  to 
the  one  with  a  lower  tractive  effort  and  higher  speed  having 
the  same  horsepower  rating,  but  he  should  even  be  cautious 
in  buying  a  locomotive  with  a  high  horsepower  rating  if  such 
a  rating  is  obtained  on  account  of  high  speed. 

The  reason  for  this  is  that  a  high  horsepower  rating  means 
increased  power  consumption  without  increasing  the  amount 
of  work  done  by  the  locomotive,  if  the  high  rating  is  obtained 
through  high  speed.  This  may  be  made  clearer  by  stating, 
for  instance,  that  a  15-hp.  locomotive  with  a  rated  speed  of, 


232 


COAL  MINING  COSTS 


say  9  miles  per  hour,  will  take  50  per  cent  more  current  for 
starting  a  certain  train  than  a  locomotive  having  a  10-hp. 
rating  with  a  rated  speed  of  6  miles  per  hour. 

It  is  evident  from  this  how  misleading  a  mere  consideration 
of  the  horsepower  rating  of  the  locomotive  may  be.  In  the 
particular  instance  given  above,  the  purchaser  might  think 
that  he  is  getting  a  locomotive  which  will  do  50  per  cent  more 
work  when  he  buys  a  15-hp.  locomotive  instead  of  a  10  hp., 
while  actually  he  would  not  be  able  to  pull  any  more  with 
such  a  machine  and  would  at  the  same  time  pay  for  his  mis- 
take in  higher  current  consumption.  As  a  matter  of  course, 


10 


15 


40 


20          25          30         35 
Radius  of  Curve,  Feet 

FIG.  3. — Curves  showing  the  relation  between  wheel  base  and  radius 

of  curve. 

there  are  certain  limitations  in  speed,  below  which  it  would 
not  be  advisable  to  go,  because  if  this  is  chosen  below  certain 
limits,  the  work  done  by  the  locomotive  would  be  reduced. 

In  view  of  the  numerous  possibilities  for  being  misled,  it 
seems  advisable  for  the  buyer  of  a  mining  locomotive  to 
request  from  the  manufacturer  the  following  information 
regarding  the  rating: 

1.  Weight  of  locomotive. 

2.  Will  the  motor  be  able  to  slip  the  wheels  of  the  locomo- 
tive at  all  conditions  of  rail? 

(These  two  questions  will  definitely  determine  the  maximum 
tractive  effort  which  the  locomotive  is  able  to  exert.) 


HAULAGE  COSTS  233 

3.  What  is  the  one-hour  rating  of  the  locomotive  according 
to  the  standardization  rules  of  the  A.I.E.E.  bases  on  a  tem- 
perature rise  of  75  deg.  C.  on  stand  test  of  the  motors? 

4.  What  are  the  tractive  effort  and  speed  of  the  locomotive 
at  the  one-hour  rating? 

5.  What  is  the  continuous  ampere  rating  of  the  motors  on 
stand  test  at  one-half  and  three-fourths  of  the  rated  voltage, 
and  what  tractive  efforts  correspond  to  these  ratings;  all  bases 
on  75  deg.  C.  temperature  rise  on  the  stand  according  to  the 
standardization  rules  of  the  A.I.E.E.  ? 

By  securing  the  above  information  and  comparing  same  for 
the  various  locomotives  in  the  market,  the  purchaser  will  be 
in  a  position  to  know  what  he  is  actually  buying.  Without 
it,  he  is  liable  to  purchase  almost  anything  without  knowing 
just  exactly  what  he  is  getting. 

It  may  be  advisable  to  cite  here  one  of  the  many  cases 
where  purchasers  have  been  misled  by  not  taking  a  little  time 
in  getting  the  above  information.  At  a  certain  mine,  tests 
were  made  on  three  locomotives  of  different  manufacture.  One 
was  a  15%-ton  machine  with  motors  rated  at  185  hp.  total, 
one  was  a  17-ton  locomotive  with  motors  rated  at  210  hp.  total. 
After  an  all  day  run,  in  which  the  number  of  car-miles  hauled 
were  practically  the  same,  the  temperature  rises  on  the  motors 
were  as  follows : 
15!/^-ton  locomotive,  rated  185  hp.,  field  52  deg.  C.,  armature 

63  deg.  C. 
17-ton  locomotive,  rated  200  hp.,  field  56  deg.   C.,  armature 

65  deg.  C. 

16-ton  locomotive,  rated  210  hp.,  field  86  deg.  C.,  armature 
85  deg.  C. 

In  other  words,  the  locomotive  with  the  smallest  horse- 
power rating  in  this  case  proved  to  be  the  best  of  the  three  in 
actual  service,  while  the  machine  with  the  highest  horsepower 
rating  not  only  showed  up  to  be  the  poorest  of  the  three,  but 
even  had  temperature  rises  exceeding  safe  limits.  This  would 
mean  a  short  life  for  the  motor  insulation  and  windings. 

Number  and  size  motors  required. — In  order  to  determine 
the  proper  equipment  for  a  mine  locomotive  it  is  necessary  to 
have  the  following  information : 


234  COAL  MINING  COSTS 

Plan  and  profile  of  the  road. 
Number  of  cars  to  be  handled  per  trip. 
Number  of  cars  to  be  handled  per  hour. 
X     Weight  of  empty  cars. 
»  Length  of  cars. 
Weight  of  load. 
Frictional  resistance  of  cars. 

Time  of  layover,  including  switching  and  making  up  trip. 
Voltage  of  circuit. 
Gage  of  track. 
Weight  of  rail. 

Eadius  and  length  of  minimum  curve. 
Spread  of  track  on  minimum  curve. 
Limiting  dimensions  which  locomotive  can  have. 
Position  and  range  of  trolley  wire. 

It  is  seldom  that  all  of  the  above  information  can  be 
obtained,  and  in  many  cases  it  is  necessary  to  make  certain 
assumptions  to  supply  the  missing  data.  This  can  only  be 
done  by  one  having  considerable  experience  in  working  out 
mining  problems. 

Motors  for  mine  locomotives  are  rated  on  the  one-hour  basis 
with  a  75-deg.  C.  rise  in  temperature.  This  rating  does  not 
indicate  the  capacity  of  the  motor  for  all-day  service,  and  is 
not  used  in  determining  its  ability  to  meet  with  a  certain  set 
of  conditions. 

The  capacity  of  a  motor  for  all-day  service  depends  upon 
the  temperature  which  the  windings  will  attain.  This  in  turn 
depends  upon  the  average  heating  value  of  the  current.  Since 
the  heat  generated  by  an  electric  current  is  proportional  to 
the  square  of  the  current  value,  the  average  heating  for  all- 
day  service  must  depend  upon  the  square  root  of  the  mean 
square  of  the  current. 

Two  motors  may  have  the  same  one-hour  rating,  but  one 
may  have  a  much  larger  continuous  capacity  than  the  other, 
due  to  better  design  and  the  proper  distribution  of  the  losses. 
A  poorly  ventilated  motor  will  in  some  cases  have  hot  spots, 
which  will  lower  the  capacity  of  the  machine.  This  is  due  to 
the  fact  that,  in  order  to  keep  these  spots  within  a  safe  tem- 
perature rise,  the  average  temperature  of  the  windings  must 
be  kept  much  lower  than  would  be  necessary  if  such  spots 
were  eliminated  by  proper  design. 

That  the  real  capacity  of  a  motor  is  its  continuous  capacity 


HAULAGE  COSTS  235 

for  all-day  service  and  not  the  rating  for  one  hour  is  apparently 
not  generally  appreciated  among  mine  operators.  The  one- 
hour  rating  depends  largely  upon  the  thermal  capacity  of  the 
motor,  while  the  continuous  rating  depends  on  the  ventilation, 
distribution  of  the  losses  and  the  capacity  of  the  machine  to 
radiate  heat. 

The  one-hour  capacity  is  not  a  fair  rating  of  a  motor  for 
the  foregoing  reason  and  also  because  the  speed  of  the  motor- 
is  not  taken  into  account.  A  fairer  way  would  be  to  rate  the 
machine  on  the  pounds  tractive  effort  at  the  wheels,  irrespective 
of  the  speed,  provided  it  is  not  considered  essential  for  com- 
mercial reasons  to  capitalize  the  increased  horsepower  ratings 
due  to  increase  in  speed. 

If  the  length  of  haul,  the  grade,  curve,  running  time  and 
time  of  layover  are  known,  the  current  for  each  part  of  the 
run  can  be  computed.  In  most  main-haulage  cases  the  locomo- 
tive will  have  a  definite  cycle  to  go  through,  this  cycle  being 
repeated  throughout  the  working  day.  If  the  square  root  of 
the  mean  square  current  for  one  cycle  can  be  found,  this  will, 
of  course,  determine  the  suitability  of  the  motor  selected  for 
the  all-day  service  as  regards  heating  capacity. 

To  illustrate  the  working  out  of  the  above  principles,  the 
following  conditions  may  be  assumed  to  exist  at  a  mine  which 
desires  to  install  electric  haulage: 

Locomotive  required 1 

Profile  as  follows: 

1300  ft.  2  per  cent  grade  against  load. 

1400  ft.  1  per  cent  grade  against  load. 

2200  ft.  level. 

Number  of  cars  to  be  handled  per  trip 20 

Number  of  cars  to  be  handled  per  hour 50 

Weight  of  empty  car 2000  Ib. 

Weight  of  load 4500  Ib. 

Total  weight  of  loaded  car 6500  Ib. 

Frictional  resistance  of  cars 30  Ib.  per  ton 

Time  of  layover,  including  switching  and  making  up  trip.  5  min.  each  end 

Voltage  of  circuit 250 

Gage  of  track 36  in. 

Weight  of  rail 30  Ib.  per  yd. 

Radius  of  minimum  curve 25  ft. 

Length  of  minimum  curve 20  ft. 

Spread  of  gage  on  curve \  in. 

Limiting  dimensions  of  locomotive,  5  ft.  wide,  4  ft.  high. 

Trolley  wire  6  in.  outside  of  rail;  height  above  rail,  4  ft.  6  in.  to  6  ft. 


236 


COAL  MINING  COSTS 


The  total  weight  of  the  trip  will  be  65  tons. 

The  limiting  condition  in  regard  to  weight  is,  of  course, 
on  the  2  per  cent  grade.  The  weight  of  the  locomotive  will  be 
found  as  follows: 

30  X  65  +  20  X  2  X  65  +  20  X  2  X  W  =  400    W 

W  =  12.6  tons  if  cast-iron  wheels  are  used ; 
W  =  9.88  tons  if  steel  wheels  are  used. 

It  would,  therefore,  be  necessary  to  use  a  13-ton  locomotive 
with  cast-iron  wheels  or  a  10-ton  locomotive  with  steel  wheels. 
Steel  wheels  should  be  used  unless  the  customer  specifies  cast- 
iron.  A  locomotive  to  negotiate  a  25-ft.  curve  should  have  a 
wheel  base  not  more  than  55  in.  with  33-in.  wheels  or  65  in. 
with  30-in.  wheels.  With  motors  tandem  hung  no  trouble  will 
be  experienced  in  keeping  below  55  in.  or  65  in.  for  a  10-ton 
locomotive.  See  Fig.  3. 


50 


Speed  M.P.H. 

01  O  W 

40-  Hf.  MINE  MO  TOK 
•Continuous  Capacity,  64  Amperes  c 

t+ISt 
'  ZOC 

IVo/t 
S 

'/* 

4000 

"I 

8000  £ 

1 

1000 

\ 

(& 

/ 

\ 

J 

r 

\ 

\ 

at 

's 

^ 

^/ 

( 

2& 

/ 

f~£ 

^J/ 

£L£t 

i. 

— 

—      — 

/ 

/ 

£00 


250 


100  150 

Amperes 

FIG.  4. — Characteristic  curves  for  a  40  hp.  mine  locomotive  motor. 

The  number  of  cars  to  be  handled  per  hour  being  50,  the 
trips  per  hour  will  be  50  -j-  20  =  2y2.  The  total  time  per  trip, 
including  layover  at  each  end,  will  therefore  be  60  -7-  2y2  — 
24  min.  Allowing  5  min.  layover  at  each  end  will  make  the 
actual  running  time  14  min. 

For  a  locomotive  of  a  given  weight  there  are,  as  a  rule,  two 
or  more  motors  to  choose  from.  For  a  10-ton  machine  these 
motors  range  from  40  to  50  hp.  in  capacity,  although  larger 
ones  are  sometimes  required  for  special  cases.  A  40-hp.  motor 


HAULAGE  COSTS 


237 


is  selected  for  the  first  trial.  The  locomotive  will  have  30-in. 
wheels  with  a  gear  ratio  of  4.78  to  1.  The  highest  gear  reduc- 
tion is  always  selected  unless  a  greater  speed  is  required  and 
can  be  obtained  without  overloading  the  motors.  The  charac- 
teristics of  this  motor  are  shown  by  the  curves  in  Fig.  4. 

This  curve  is  made  from  an  actual  test,  and  the  tractive 
effort  given  includes  gear  losses,  so  that  to  obtain  the  drawbar 
pull  only  the  locomotive  friction  should  be  deducted.  A  table 
of  data  covering  the  case,  such  as  shown  in  Table  I  or  II,  should 
be  prepared,  the  values  inserted  being  calculated  from  the 
motor  curves  and  weights  to  be  handled.  Since  the  curves 
give  values  for  one  motor,  the  locomotive  and  trailing  weight 
should  be  divided  by  two  to  give  the  weight  each  motor  will 
be  required  to  handle. 

TABLE  I 


Speed, 
Miles 
per 

Amp. 

Total 
Tr. 
Eff. 

Loco. 
Res. 

Train 
Res. 

Grade 
Res. 

Dis- 
tance, 
Feet 

Time, 
Sec. 

Amp.2 

Amp.2x 
Time 

Hr. 

- 

8.5 

87 

1050 

75 

975 

0 

2200 

177 

7,569 

1,340,000 

6.5 

165 

2550 

75 

975 

1500 

1300 

136 

27,225 

3,700,000 

7.2 

125 

1800 

75 

975 

750 

1400 

133 

15,625 

2,080,000 

Returning  with  Empty  Trip 

10 

25 

75 

75 

300 

-300 

1400 

96 

625 

60,000 

10 

0 

-225 

75 

300 

-600 

1300 

89 

10 

45 

375 

75 

300 

0 

2200 

150 

2,025 

303,000 

Amp.2  X  time  = 

7,483,000 

Plus  10  per  cent  for  accelerating,  switching,  etc  .... 

748,300 

Total  amp  2  X  time            

8,231,300 

Total  running  time,  sec  

781 

Total  time  at  both  ends,  sec  

659 

Total  time  including  lavover.  sec  .  . 

1.440 

8,231,300-^- 1440  =  5700  =  mean  squared  current. 

The  square  root  of  5700  =  75 . 5  =  square  root  of  mean  square  current. 

Capacity  of  40-hp.  motor  is  60  amp. 


For  the  above  project  the  weight  of  locomotive  is  five  tons 
per  motor;  the  loaded  trailing  weight,  32%  tons  per  motor, 
and  the  light  trailing  weight,  10  tons  per  motor.  Assume  that 
the  locomotive  starts  with  a  load  on  a  level  track  and  runs 
2200  ft.,  when  it  encounters  a  2  per  cent  grade.  After  ascend- 


238  COAL  MINING  COSTS 

ing  this  grade  for  1300  ft.  the  grade  changes  to  1  per  cent  for 
1400  ft.  The  return  trip  will  be  with  empty  cars. 

Starting  with  the  loaded  trip  on  the  level,  the  locomotive 
resistance  per  motor  will  be  5  X  15  =  75  Ib.  This  value  is 
placed  under  "locomotive  resistance"  in  the  table.  The  train 
resistance  will  be  32.5  X  30  —  975  Ib.  The  grade  resistance 
will  be  zero.  The  total  tractive  effort  will  be  1050  Ib. 

Consulting  the  motor  curves  of  Fig.  4  the  current  for  a 
tractive  effort  of  1050  Ib.  is  87  amp.  and  the  speed  8.5  miles 
per  hour.  The  time  to  cover  2200  ft.  at  8.5  miles  per  hour  will 
be  177  sec.  The  amperes  squared  will  be  7569  and  the  amperes 
squared  multiplied  by  time  will  be  1,340,000.  These  values 
should  be  recorded  in  their  proper  place  in  the  table. 

When  the  2  per  cent  grade  is  reached,  the  train  and  loco- 
motive resistance  will  remain  the  same,  while  the  grade  resist- 
ance will  be  40  X  (32.5  +  5)  =1500  Ib.  The  total  tractive 
effort  will  be  2550  Ib.,  which  corresponds  to  a  motor  current 
of  165  amp.  and  a  speed  of  6.5  miles  per  hour.  At  this  speed 
it  will  require  143  sec.  to  travel  1300  ft.  By  the  same  process 
the  values  for  the  1  per  cent  grade  are  calculated  and  filled  in 
the  table. 

On  the  return  trip  with  the  empty  cars  the  locomotive 
resistance  will  be  the  same,  the  train  resistance  300  Ib.  for 
10  tons,  and  the  down-grade  resistance  —  300  Ib.  for  the  1  per 
cent  and  —  600  Ib.  for  the  2  per  cent  grade.  It  will  be  noted 
that,  running  down  the  2  per  cent  grade,  the  net  tractive 
effort  is  —  225  Ib.,  which  means  that  the  brakes  must  be 
applied  in  descending  this  grade. 

On  the  1  per  cent  grade  and  on  the  level  the  speeds  shown 
on  the  curve  of  Fig.  4  are  too  high  for  most  mines  unless  the 
track  is  in  good  shape.  It  is  likely  that  the  operator  will  not 
care  to  run  faster  than  10  miles  per  hour,  which  he  can  do  by 
operating  the  motors  in  series  on  low  notches,  or  by  cutting 
off  the  power  and  coasting  before  the  speed  becomes  too  high. 

The  total  running  time  is  781  sec.,  or  13  min.  1  sec.  The 
actual  running  time  will  be  close  to  14  min.,  due  to  time 
taken  to  start  and  stop  the  trip  and  for  possible  slowing  down 
at  cross-overs.  As  the  total  time  for  a  round  trip  is  24  min., 
a  layover  of  five  minutes  is  obtained  at  each  end. 

The  product  of  the  total  current  squared  by  the  time  is 


HAULAGE  COSTS 


239 


7,483,000.  To  this,  10  per  cent  should  be  added  to  allow  for 
acceleration  and  switching  when  making  up  trips,  making  a 
total  of  8,231,300.  The  total  time  for  making  a  round  trip, 
including  layover,  is  1440  sec.  Dividing  8,231,300  by  1440  = 
5700  as  the  mean  square  of  the  current.  The  square  root  of 
5700  is  75.5  amp.,  which  is  the  square  root  of  the  mean  square 
current  for  one  trip  or  cycle. 

The  continuous  capacity  of  the  motor  is  68  amp.  at  150 
volts  and  64  amp.  at  200  volts.  The  class  of  service  is  such 
that  the  average  voltage  applied  to  the  motor  will  be  near 
200,  so  that  the  rating  of  the  motor  is  about  65  amp.,  which 
shows  that  it  is  not  of  sufficient  capacity  for  the  service. 

TABLE  II 


Speed, 
Miles 
per 
Hr. 

Amp. 

Total 
Tr. 

Eff. 

Loco. 
Res. 

Train 
Res. 

Grade 
Res. 

Dis- 
tance, 
Feet 

Time, 
Sec. 

Amp.2 

Amp.2x 
Time 

10.1 
7.6 
8.5 

100 
187 
146 

1050 
2550 
1800 

75 
75 
75 

975 
975 
975 

0 
1500 
750 

2200 
1300 
1400 

149 
117 
113 

10,000 
34,970 
21,316 

1,490,000 
4,090,000 
2,410,000 

Returning  with  Empty  Trip 


75           75           300 
25           75           300 
75           75           300 

accelerating,  switching 

-300 
-600 
0 

,  etc.  ... 

1400 
1300 
2200 

96 
89 
150 

1,225 
3,136 

117,500 
470,400 

8,577,900 
857,790 

9,435,690 

714 
726 

layover,  sec  .  . 

1.440 

Amp. 2X  time: 


Total  amp.2  X  time 
Total  running  time,  sec .  . 
Total  time  at  both  ends,  sec 


9,435,690  -5-  1440  =  6550  =  mean  squared  current. 

The  square  root  of  6550  =  81  =  square  root  of  mean  square  current. 

Capacity  of  50  hp.  motor  is  80  amp.  at  150  v. 


Care  should  be  taken  that  a  motor  is  not  selected  in  which 
the  commutating  limit  is  exceeded  when  the  wheels  are  slipped 
while  using  sand.  The  motor  curves  are  generally  stopped 
at  the  commutating  limit,  although  with  the  modern  com- 
mutating-pole  motor  it  is  rather  difficult  to  find  the  commutat- 
ing limit. 


240 


COAL  MINING  COSTS 


A  larger  motor  should  be  selected,  and  Table  II  shows  the 
results  of  the  calculation  using  two  50-hp.  motors.  The  curves 
of  Fig.  5  show  the  characteristics  of  this  motor.  The  total 
calculated  running  time  is  714  sec.,  or  11  min.  54  sec.  The 
square  root  of  the  mean  square  current  is  found  to  be  SO  amp. 
The  capacity  of  the  motor  at  200  volts  is  78  amp.  and  82  amp. 
at  150  volts,  so  that  this  motor  will  be  of  just  about  the  proper 
capacity  to  meet  the  conditions.  The  actual  running  time  will 
be  about  12  to  14  min.,  allowing  5  to  6  min.  at  each  end  for 


16 


50-HP  MINE  MOTOR 
Continuous  CapacHy,  82 Amperes  erf-  ISO  Vo 
r»    y  78     r»        »  £00 . 


5000 


4000 


3000 


2000 


1000 


0      40     60      120    160    200  Z40  280   320  360 
Amperes 

FIG.  5. — Characteristic  curves  for  a  50-hp.  locomotive  motor. 

switching  and  making-up  trips.  The  higher  root  mean  square 
current  obtained  for  the  50-hp.  motor  is  due  to  the  fact  that 
it  is  a  higher-speed  machine  than  the  40-hp.  This  shows  the 
importance  of  having  a  low-speed  motor. 

It  is  not  safe  to  figure  on  a  short  layover,  as  in  many 
cases  the  average  line  voltage  is  much  less  than  500  or  250  volts, 
which  means  that  the  speed  will  be  less  than  is  figured  on. 
A  low  line  voltage  signifies  that  a  given  current  will  be  re- 
quired for  a  much  longer  time  than  with  the  normal  voltage, 
which  in  turn  means  additional  heating.  Where  the  voltage 
is  likely  to  be  poor,  a  margin  should  be  allowed  in  the  motor 


HAULAGE  COSTS 


241 


capacity,  since  the  value  of  the  square  root  of  mean  square 
current  will  be  greater  than  that  calculated. 

The  conditions  outlined  above  in  regard  to  profile  are 
typical  of  what  may  be  expected  in  mines.  Sometimes  it  is 
possible  to  lay  out  a  mine  so  that  all  of  the  main  haulage  will 
be  with  the  grades ;  other  mines  will  have  mixed  grades — that 
is,  some  inclinations  in  favor  of  the  load  and  some  against  it ; 
while  too  often  a  mine  will  be  found  with  little  or  no  level 
track  and  all  grades  against  the  load.  If  the  profile  instead 
of  that  given  above  were  2  per  cent  against  load  300  ft.,  level 
track  2300  ft.,  1  per  cent  in  favor  of  load  2300  ft.,  the  same 
weight  of  locomotive  would  be  required,  but  the  heating  would 
be  much  less.  Table  III  shows  that  the  root  mean  square 
current  is  only  51.7  amp.,  which  gives  a  large  margin  of  safety 
when  using  the  40-hp.  motor. 


TABLE  III 


Speed, 
Miles 
per 
Hr. 

Amp. 

Total 
Tr. 

Eff. 

Loco. 
Res. 

Train 
Res. 

Grade 
Res. 

Dis- 
tance, 
Ft. 

Time, 
Sec. 

Amp.2 

Amp.2x 
Time 

6.5 
10 
8.5 

165 
40 

87 

2550 
300 
1050 

75 
75 
75 

975 
975 
975 

1500 
-750 
0 

300 
2300 
2300 

31.5 
157 
184.5 

27,225 
1,600 
7,569 

870,000 
251,200 
1,395,000 

10 
10 
10 


Returning  with  Empty  Trip 


75           75 
75           75 
25           75 

iccelerating,  g 

300                0          2300        157 
300            300          2300        157 
300         -600            300          20 

witching,  etc 

2,025 
4,225 

318,000 
664,000 

3,498,200 
349  820 

3  848  020 

ec  

707 

ids,  sec  

733 

layover,  sec. 

1.440 

Amp.2  X  time 


Total  amp.2  X  time 
Total  running  time,  sec 
Total  time  at  both  ends,  sec 


3,848,020  -=- 1440  =  2670  =  mean  squared  current. 

The  square  root  of  2670  =  51.7  amp.  =  square  root  of  mean  square  current. 


On  the  other  hand,  the  grade  may  be  2  per  cent  for  the 
entire  distance  against  the  load.  In  this  case  the  root  mean 
square  current  would  be  about  105  amp.,  corresponding  to  a 
motor  having  an  hour  rating  of  75  to  80  hp.  As  it  would  not 
be  practical  to  put  two  such  large  motors  on  a  well  designed 


242  COAL  MINING  COSTS 

10-ton  locomotive,  it  would  be  necessary  to  go  to  a  12-  or  13-ton 
machine. 

In  another  computation  carried  out  along  somewhat  dif- 
ferent lines,  it  will  be  assumed  that  it  is  desired  to  haul  coal 
from  a  siding  at  the  bottom  of  the  slope,  in  trips  of  30  cars. 
The  empty  cars  each  weigh  1  ton  and  have  a  capacity  of 
1.5  tons.  This  will  make  the  total  weight  of  the  loaded  trip 
30  (1  -|-  1.5)  =75  tons.  The  accompanying  profile  of  the  road, 
Fig.  6,  extending  from  the  siding  at  the  slope  bottom  in  the 
mine  to  the  tipple  where  the  coal  is  loaded  into  the  railroad 
cars,  shows  the  length  and  grade  of  each  portion  of  the  track 
and  the  weight  of  rail  in  use. 

It  is  desired  to  know  the  size  of  electric  motor  that  will  be 
required  for  this  haul;  also  the  weight  of  rail  and  style  of 
rail  bond  that  should  be  used,  the  size  of  trolley  wire  required 
for  the  transmission  of  the  power  from  the  generator  to  the 
mine,  and  the  required  horsepower  of  the  generator  and  boiler. 

The  first  step  is  to  estimate  the  weight  of  locomotive  re- 
quired to  haul  a  loaded  trip  of  75  tons  up  a  3-per  cent  grade, 
under  ordinary  mining  conditions.  It  is  not  safe  to  estimate 
on  the  track  resistance  as  being  less  than  50  Ib.  per  ton.  To 
this  must  be  added  20  Ib.  per  ton  for  each  per  cent  of  grade 
or,  in  this  case,  3  X  20  =  60  Ib.  grade  resistance,  which  makes 
the  total  resistance  50  -f-  60  =  110  Ib.  per  ton  of  total  moving 
load  including  the  locomotive. 

The  coefficient  of  adhesion  of  the  wheels  to  the  rails  will 
be  estimated  at  0.16,  which  makes  the  tractive  effort  that  the 
locomotive  can  exert  to  move  itself  and  the  loaded  trip  0.16 
X  2000  =  320  Ib.  per  ton.  Then,  since  this  tractive  effort  of 
the  locomotive  must  be  equal  to  the  total  resistance  of  the 
entire  moving  load  including  the  locomotive, 


and 

^         110TF,       110X75 

UQ  =  ~2I6~ 


Ar. 

=Sa    4° 


This  shows  conclusively  that,  under  the  adverse  conditions 
common  in  coal  mining,  a  40-ton  locomotive  would  be  required 
to  haul  a  loaded  trip  of  75  tons  up  a  3-per  cent  slope.  This  of 
course  provides  for  the  worst  conditions  that  are  liable  to 


HAULAGE  COSTS 


243 


8.   8 


£ 


&'£ 


244  COAL  MINING  COSTS 

exist,  in  respect  to  both  the  track  and  rolling  stock  in  the 
mine. 

Before  going  further,  it  will  be  well  to  estimate  the  weight 
of  locomotive  that  will  be  required  to  haul  the  same  loaded 
trip  on  the  1%-per  cent  grades  outside  of  the  mine.  In  this 
case,  the  track  resistance,  as  before,  is  50  Ib.  per  ton,  but  the 
grade  resistance  is  1.5  X  20  =  30  Ib.  per  ton,  which  makes  the 
total  resistance  50  -f-  30  =  80  Ib.  per  ton.  This  gives  for  the 
required  weight  of  the  locomotive : 


80X75 
320-80 


TTr  O\J  /\  I  fj  c\f     , 

Wm = ^^ — ™  =  25  tons. 


The  same  weight  locomotive  will  haul  practically  two-thirds 
of  the  number  of  cars  up  the  slope  that  it  can  handle  on  the 
outside  grades,  as  shown  by  transposing  the  formula  pre- 
viously given  and  finding  the  load  that  a  locomotive  can  be 
expected  to  haul  regularly  up  grades  of  1%  and  3  per  cent, 
respectively,  under  the  conditions  named,  thus, 

On  a  3-per  cent  grade: 

tractive  effort  of  locomotive,  320  Ib.  per  ton; 
track  and  grade  resistance,       80  Ib.  per  ton. 

qon  _  1  in 

Loaded  trip,  Wt=  Wm=  1.9TFW. 


On  a  IJ-per  cent  grade: 

tractive  effort  of  locomotive,  320  Ib.  per  ton; 
track  and  grade  resistance,       80  Ib.  per  ton. 

Loaded  trip,  Wt  =  32°8~8°  Wm  =  3Wm. 

For  this  reason,  it  might  be  well  to  provide  a  siding  at  the 
top  of  the  slope  that  will  hold  20  or  40  cars  as  desired,  and 
make  three  trips  on  the  slope  for  every  two  trips  to  the  tipple, 
using  a  25-ton  locomotive  for  the  entire  work. 

So  far,  we  have  only  considered  relative  weights  of  the 
locomotive  and  the  load  it  can  be  expected  to  handle  regularly 
and  satisfactorily  on  different  grades,  under  the  worst  con- 


HAULAGE  COSTS  245 

ditions  that  are  liable  to  exist  or  arise  in  mining  practice.  It 
is  important  to  remember  that  having  assumed  a  coefficient 
expressing  the  adhesion  of  the  wheels  to  the  rails  ft  is  the 
weight  of  the  locomotive  that  determines  the  tractive  effort  the 
machine  can  exert,  regardless  of  the  power  of  the  motors  or 
engines  with  which  it  is  equipped.  It  is  a  fatal  mistake  to 
equip  a  locomotive  with  more  power  than  its  weight  will 
permit  it  to  utilize. 

Having  decided  on  the  weight  of  the  locomotive  necessary 
to  haul  the  desired  number  of  loaded  cars  over  the  given 
grades,  the  next  step  is  to  determine  the  power  of  the  motors 
that  will  be  required  to  produce  a  given  speed  of  haul,  say 
from  6  to  8  mi.  per  hr.  As  a  basis  of  this  calculation,  it  is 
important  to  observe  that  the  effective  power,  in  foot-pounds 
per  minute,  is  equal  to  the  tractive  effort  in  pounds,  multiplied 
by  the  speed  in  feet  per  minute,  as  expressed  by  the  formula. 

Power  =  tractive  effort  X  speed. 

(ft.-lb.p.m.)  (lb.)  (ft.p.m.) 

Now,  estimating  the  horsepower  required  per  ton  on  drivers, 
per  mile-hour,  assuming  a  tractive  coefficient  c  =  0.16,  we  have: 
Tractive  effort  (per  ton)  =0.16X2000  =  320  lb. 
Speed  (per  mi.-hr.)  =  5280  ^  60  =  88  ft.p.m. 

QOQ  \/  CO 

Horse-power  (per  ton-mi. -hr.)-. —  =0.85  h.p. 

»  ooUUU 

This  is  the  effective  horsepower  per  ton  mile-hour,  or  the 
power  that  must  be  available  to  produce  a  speed  of  1  mi.  per  hr., 
for  each  ton  of  weight  resting  on  the  drivers.  Assuming  an 
efficiency  of  85  per  cent,  the  input  to  the  motors  should  be 
1  hp.  per  ton-mi.-hr. 

Using  a  25-ton  locomotive  for  this  haulage  and  assuming 
that  the  entire  weight  of  the  locomotive  is  on  the  drivers,  the 
actual  horsepower  of  the  motors  required  for  a  speed  of  6  mi. 
per  hr.  will  be  25  X  6  =  150  hp.,  which  corresponds  to  a  wat- 
tage of  150  +  746  =  111,900  watts  delivered  to  the  motors. 

It  is  interesting  to  note  here  that  the  effective  wattage  per 
mile-hour,  in  electric  mine  haulage,  is  practically  twice  the 
tractive  effort  of  the  locomotive,  in  pounds.  This  is  shown 
by  the  following  simple  calculation,  since  a  speed  of  1  mi.  per 


246  COAL  MINING  COSTS 

hr.  is  equal  to  88  ft.  per  min.  and  1  hp.  is  equivalent  to  746 
watts : 

/OO  y  rp  p  \ 

Effective  watts  (per  mi.-hr.)74Q=  (    33QQQ    j=say  2  T-E- 

Under  the  assumed  conditions,  it  was  previously  estimated 
that  the  total  tractive  effort  when  hauling  up  a  3-per  cent 
grade  is  110  Ib.  per  ton  of  moving  load,  which  makes  the 
effective  wattage  required  in  that  case  2  X  HO  —  220  watts 
per  ton-mi.-hr.  In  like  manner,  it  was  estimated  that  the  total 
tractive  effort  when  hauling  up  a  1%-per  cent  grade  was  80  Ib. 
per  ton,  which  makes  the  effective  wattage  required  to  haul 
up  that  grade  2  X  80  =  160  watts  per  ton-mi.-hr. 

In  electric  mine  haulage,  it  is  customary  to  estimate  on 
hauling  at  a  speed  of,  say  from  6  to  8  mi.  per  hr.  Taking 
the  actual  power  of  the  motors  at  1  hp.  per  ton-mi.-hr.,  to  give 
an  input  of  111,900  watts  on  a  500-volt  circuit  will  require  a 
current  of  111,900  -r-  500  =  say  224  amp. 

The  weight  of  rail  in  pounds  per  yard,  in  locomotive  mine 
haulage,  may  be  taken  as  twide  the  weight  of  the  locomotive 
in  tons,  or,  in  this  case,  2  X  25  =  50  Ib.  per  yd.  In  estimating 
the  size  of  wire  required  to  transmit  this  current  from  the 
generator  to  the  siding  at  the  foot  of  the  slope  in  the  mine, 
we  will  assume  a  rail-return,  using  50-lb.  rails  for  the  track, 
bonded  with  compressed-terminal  bonds,  24  in.  long.  The 
length  of  track  from  the  end  of  the  siding  at  the  foot  of  the 
slope  to  the  tipple  is  about  5400  ft.  The  length  of  trolley  wire 
extending  to  the  power  plant  will  be  assumed  as  5600  ft. 

Referring  now  to  the  diagram  shown  in  Fig.  7,  first  find 
the  resistance  of  5200  ft.  of  the  two  50-lb.  rails  in  the  track- 
return,  including  the  resistance  of  the  bonds,  assuming  a  rail 
to  copper  ratio  of  1  : 11  and  rails  30  ft.  long,  making  5400 
-4-  30  =  say  180  bonds  in  a  single  length  of  rail.  Following 
the  horizontal  line  marked  50  on  the  left  of  the  diagram,  to 
its  intersection  with  the  curved  line  indicating  a  rail-to-copper 
ratio  of  1  : 11 ;  and,  from  that  point,  following  the  vertical  line 
to  the  scale  at  the  bottom  of  the  diagram,  we  find  a  resistance 
of  18  microhms  per  foot  of  rail  or  30  X  18  =  540  microhms  per 
rail.  In  like  manner,  for  a  24-in.,  4-0  bond,  follow  the  vertical 
line  marked  24  on  the  top  scale  down  to  its  intersection  with 


HAULAGE  COSTS 


247 


the  diagonal  4-0;  and,  from  this  point,  follow  the  horizontal 
line  to  the  right-hand  scale,  which  shows  a  bond-resistance  of 
100  microhms. 

The  total  resistance  for  180  bonded  rails  is,  therefore,  180 
(540  +  100)  =115,200  microhms,  or  0.1152  ohm.  The  resist- 
ance for  a  two-rail  return  is  one-half  this  amount  or  0.0576 
ohm.  The  voltage  absorbed  in  the  rail-return  is  E  =  CR  = 
224  X  0.0576  =  12.9  volts.  The  line  drop  for  5600  ft.  of  T.  B., 
4-0,  copper  wire,  carrying  a  current  of  224  amp.,  as  taken.  «- 
from  wire  tables,  is  5.6  (224  X  0.0489)  =  61.3  volts.  The  dif-K 


Length  of  Bond,  Inches 
12       16      20      24       26 


32      36 


<b       6 


24      26 


10       12        14       16       16       20       22 

"Rail  Resistance,  Microhms  per  Ft. 
FIG.  7. — Resistance  diagram  for  bonded  rails. 

ference  of  potential  at  the  generator  is  therefore  61.3  -f- 12.9 
-f  500  =  574  volts. 

Finally,  assuming  an  efficiency  of  90  per  cent,  the  power 
required  to  run  the  generator  is : 
Input  to  generator, 

(574  X  224)  -f-  (0.90  X  746)  =  say  190  hp. 
Boiler  horsepower,  efficiency  of  engine  being  90  per  cent. 
190  -=-  0.90  =  say  210  hp. 

The  narrow  track  gages  now  used  for  mine  locomotives 
limit  the  axial  length  of  the  motor  to  a  few  inches.    The  desira- 


248 


COAL  MINING  COSTS 


bility  of  obtaining  the  maximum  tractive  effort  from  a  given 
machine  by  using  the  smallest  possible  pinion,  limits  the  gear- 
center  distance,  and  therefore  the  horizontal  width  of  the  motor, 
while  restricted  head  room  limits  the  height  of  the  motor. 

In  Figs.  8  and  9  are  shown  two  curves  taken  from  a  paper 
by  G-.  M.  Eaton,  which  demonstrates  the  fact  that  these  con- 


1900      »902       1904       1906       I9Q6       1910 
Years 


1912      1914 


FIG.    8. — Maximum  and  minimum  height  of  mining  locomotives, 
1896  to  1914. 


1898       1900       1902      1904       1906       1908       1910       1912        1914 
Years 

FIG.  9. — Tractive  effort  of  mine  locomotives  per  inch  of  height  and 
of  track  gage. 

ditions  are  becoming  more  and  more  exacting  each  year.  One 
curve  shows  the  minimum  height  allowed  for  mining  locomo- 
tives over  a  period  extending  from  1896  to  1914,  from  which 
it  will  be  seen  that  there  has  been  a  continuous  decrease  in 
minimum  head  room.  The  other  curve  shows  the  relative 
tractive  efforts  obtained  per  inch  of  height  of  locomotive  as 


HAULAGE  COSTS  249 

well  as  that  per  inch  of  gage.  It  will  be  seen  from  these  curves 
that  there  exists  a  continuous  demand  for  increasing  the  motor 
rating,  crowded  into  a  certain  space. 

The  number  and  size  motors  used  at  different  mines  varies 
over  a  considerable  range,  the  following  being  some  typical  ex- 
amples : 

The  Peabody  Kincaid  Mine  in  Illinois  uses  13-ton  motors  on 
the  primary  haulage  and  6-ton  on  the  secondary.  The  Madison 
Coal  Corporation  at  its  No.  6  Mine  in  the  same  state  uses  three 
15-ton  motors  hauling  25  to  30-car  trips  and  two  12-ton  motors 
handling  20  to  25-cars  trips.  These  motors  are  used  exclusively 
on  the  main  haulage,  there  being  four  hauls  of  7000  ft.  and 
five  hauls  of  5000  ft.  Secondary  haulage  is  done  entirely  with 
mules,  there  being  40  used  for  this  purpose.  A  comparison  of 
the  number  of  tons  handled  per  motor  at  the  mines  in  Guernsey 
County,  Ohio,  showed  this  to  vary  from  250  to  500. 

At  a  West  Virginia  mine,  of  small  capacity,  haulage  costs  per 
a  single  month  (July,  1916)  amounted  to  5.09c.  per  ton.  Room 
tracks  at  this  mine  were  laid  with  wooden  rails,  mules  were  used 
for  gathering,  and  a  motor  on  the  main  haul,  which  was  about  one 
mile  long.  The  capacity  of  the  mine  cars  was  2500  Ib. 

Another  West  Virginia  operation  handled  35,000  tons  in  one 
month  at  the  cost  of  less  than  5c.  per  ton  (September,  1916). 
This  mine  used  a  10-ton  motor  on  the  main  haul  and  a  6-ton  on 
the  secondary. 

In  the  middle  Western  fields  where  large  capacity  cars  pre- 
dominate, it  has  been  found  that  under  average  conditions  a 
5-ton  gathering  motor  will  handle  from  80  to  125  cars  per  8  hr. 
shift.  These  cars  weigh  one  ton  empty,  have  a  capacity  of  3!/2 
tons,  and  the  motor  handles  15  loads  on  a  level  track  or  six  to 
eight  on  a  fairly  stiff  grade. 

Experience  in  another  field  has  shown  that  a  6-ton  gathering 
motor  working  under  average  conditions  will  handle  80  to  90 
cars  of  2!/o  ton  capacity  in  10  hr. ;  where  the  average  haul  does 
not  exceed  2000  ft.  and  under  fairly  good  conditions,  it  will 
handle  100  to  125  cars.  Computing  on  the  basis  of  an  actual 
working  time  of  8  hr.,  to  make  allowance  for  changing  cars  and 
unavoidable  delays,  and  allowing  1%  tons  for  the  weight  of  the 
empty  car,  this  motor  is  performing  20  ton-miles  of  work  an  hour. 

In  general,  small  locomotives  are  more  satisfactory  than  large 


250  COAL  MINING  COSTS 

ones.  By  small  locomotives  are  meant  those  weighing  from  5  to 
10  tons.  There  are  a  number  of  15-ton  four-wheeled  locomotives 
in  fairly  satisfactory  service,  and  a  few  of  this  type  weighing  20 
tons  are  in  use.  Because  of  track  conditions  locomotives  weighing 
15  or  18  tons  should  have  six  wheels.  When  it  is  desirable  to 
use  machines  underground  weighing  20  tons  or  more  the  best 
results  are  obtained  by  using  two  four-wheeled  units  connected 
in  tandem. 

In  a  few  special  cases  it  is  economical  to  use  locomotives  weigh- 
ing around  30  tons,  in  which  case  it  is  advisable  to  employ  two 
units  connected  in  tandem  and  operated  from  a  250-volt  trolley 
because  of  the  difficulty  of  collecting  the  large  amount  of  power 
required  from  a  single  wire.  In  general,  when  the  size  of  the 
locomotive  exceeds  15  tons,  current  at  500  volts  should  be  used 
to  secure  good  results. 

The  Berwind- White  Coal  Mining  Co.  in  1915  placed  in 
operation  at  Windber,  Penn.,  what  was  then  the  two  largest  and 
most  powerful  single-unit  electric  mine  locomotives  in  the  world. 
These  machines  weigh  30  tons  each,  and  are  of  the  three-motor, 
six-wheeled  type,  with  equalizing  levers  to  evenly  distribute  the 
weight  upon  the  drivers.  The  side  and  end  frames  are  con- 
structed of  what  is  known  as  " armor  plate/'  solid  rolled  steel 
slabs,  the  sides  being  5-in.  thick,  40  in.  high  and  17  ft.  long. 

These  members  by  the  aid  of  angles  of  heavy  steel  casting 
are  securely  bolted  together,  forming  a  rigid  unit. 

The  motor  equipment  of  each  locomotive  consists  of  three 
115-hp.,  500-volt  ball-bearing  motors,  capable  of  developing  a 
tractive  effort  of  15,500  Ib.  at  a  speed  of  8  miles  per  hour.  In- 
corporated in  the  design  of  these  motors  are  the  essential  mechan- 
ical and  electrical  features,  which  have,  to  a  large  extent,  solved 
the  troubles  encountered  in  mining  work.  The  application  of  ball 
bearings  permits  the  use  of  gearing  of  51/*)  in.  face,  and  a  commu- 
tator 5%  in.  wide,  within  the  space  available  on  a  36-in.  track 
gage,  without  a  sacrifice  in  size  of  any  part. 

An  electric  mine  locomotive  that  has  remarkable  record  for 
length  of  service  is  at  present  operating  in  a  mine  of  the  Union 
Pacific  Coal  Co.  at  Rock  Springs,  Wyo.  This  locomotive  was 
put  into  service  in  its  present  location  27  yr.  ago.  It  has  been 
giving  continuous  service  ever  since,  records  kept  by  the  Union 
Pacific  Coal  Co.  showing  that  it  has  hauled  3,712,500  tons  of 


HAULAGE  COSTS  251 

coal  over  an  average  distance  of  1.5  miles,  making  a  total  of 
5,568,750  ton-miles. 

This  locomotive  is  a  terrapin  back  machine  built  for  500  volts 
and  having  a  speed  of  8  miles  per  hour.  It  has  a  drawbar  pull 
of  3000  Ib.  and  the  wheels  are  28  in.  in  diameter. 

Gathering  Locomotives. — When  a  locomotive  is  to  be  used 
for  gathering  it  is  difficult  to  determine  the  proper  capacity  by 
the  preceding  methods,  since  the  service  consists  largely  of  start- 
ing and  stopping  with  varying  loads.  From  experience  it  has 
been  found  that  if  the  horsepower  per  ton  of  weight  of  locomotive 
ranges  from  6  to  10,  the  motors  will  have  ample  capacity. 

This  scheme  is  much  of  a  makeshift  and  has  little  real 
reason  behind  its  use.  Locomotives  have  been  in  successful 
operation  for  years  with  but  6  hp.  per  ton  and  the  motors  were 
not  overloaded.  On  the  other  hand,  cases  sometimes  arise 
where  12  to  14  hp.  per  ton  would  be  entirely  inadequate.  A 
high  horsepower  per  ton  is  often  obtained  by  using  a  high- 
speed motor  and  either  allowing  the  locomotive  to  run  fast 
or  to  use  a  pinion  which  is  entirely  too  small  for  safe  operation. 

The  placing  of  too  large  an  equipment  on  a  locomotive  is 
a  detriment  instead  of  an  advantage.  It  is  somewhat  similar 
to  hitching  a  heavy  dray  horse  to  a  light  express  wagon.  The 
horse  cannot  work  up  to  anything  like  his  capacity  and  is  likely 
to  injure  the  wagon  in  attempting  to  perform  his  normal  work. 

Where  the  motor  is  too  large  the  speeds  will  be  high  and 
the  motor-man  is  often  tempted  to  use  an  excess  of  sand  and  to 
hold  the  brakes  on  during  acceleration  to  prevent  slipping. 
This  kind  of  operation  results  not  only  in  a  rise  in  power  con- 
sumption but  also  a  large  increase  in  mechanical  wear.  The 
saving  in  electrical  repairs  may  be  more  than  balanced  by  the 
increase  in  mechanical  repairs. 

When  reels  are  desired  they  can  be  furnished  either  of 
the  traction  or  electrical  type.  The  traction  reel  is  used  where 
steep  grades  are  encountered  and  it  is  not  desirable  to  have 
the  locomotive  enter  the  rooms.  The  capacity  of  the  motor 
for  a  traction  reel  should  be  from  4  to  8  hp.,  depending  upon 
the  conditions.  A  small  resistance  type  of  controller  should  be 
used. 

Of  electrical  reels  two  general  types  are  used — the  mechan- 
ically and  the  electrically  driven.  The  mechanically  driven  reel 


252  COAL  MINING  COSTS 

is  geared  to  one  of  the  axles  or  to  one  of  the  main  gears.  This 
reel  has  the  disadvantage  that  if  the  wheels  are  locked  by  the 
brake  when  coming  out  of  a  room  the  cable  will  be  run  over 
and  cut  in  two. 

The  electrically  driven  reel  consists  of  a  horizontal  or  ver- 
tical axle  reel  equipped  with  a  small  motor,  which  acts  as  an 
electrical  spring,  always  keeping  tension  on  the  cable.  The 
motor  is  generally  less  than  1  hp.  in  capacity  and  is  so  wound 
that  it  can  remain  connected  across  the  line  continuously  with- 
out danger  of  being  burned  out.  The  horizontal  reel  has  the 
advantages  that  it  does  not  interfere  with  access  to  the  main 
motor  and  can  be  mounted  low  down  in  front  without  increas- 
ing the  height  of  the  locomotive. 

Costs. — A  leading  manufacturer  of  electric  mine  motors 
quoted  prices,  f.o.b.  mines,  January,  1921,  as  follows:  4-ton, 
$3950 ;  same  with  gathering  reel,  $4500 ;  6-ton,  $4595 ;  and  8-ton, 
$5540,  both  the  latter  being  without  gathering  reels.  The  life  of 
of  mine  motors  may  be  estimated  at  25  yr. 

R.  V.  Norris,  in  the  transaction  of  the  A.  I.  M.  M.  E.,  Vol.,  34, 
p.  976,  gives  interesting  figures  on  the  cost  of  the  first  electric 
underground  haulage  system  installed  in  this  country,  and  it  is 
believed  the  second  in  the  world,  which  was  started  July  23,  1887, 
at  the  Short  Mountain  Colliery  of  the  Lykens  Valley  Coal  Co. 

The  plant  consists  of  two  5-ton  motors  of  32  hp.  each,  two 
65  hp.  generators  driven  by  an  old  18  X  24  in.  plain  slide-valve 
engine.  The  average  voltage  is  450,  the  amperage  from  40  to 
200,  the  average  indicated  horsepower  of  the  engine,  about  75, 
with  a  steam  consumption  of  about  75  Ib.  per  indicated  horse- 
power. The  cost  of  steam  at  the  colliery  was  8.11c.  per  1000  Ib., 
so  that  the  steam  cost  per  day  was  $4.56. 

There  were  two  main  hauls  at  this  colliery,  one  of  9500  ft.  and 
the  other  of  10,400  ft.  The  average  trip  was  made  up  of  15  cars, 
weighing  one  ton  each  and  carrying  2.25  tons  of  coal  or  3  tons  of 
rock.  The  following  is  the  work  done  with  this  system  in  the 
year  1901 : 

The  total  cost  for  labor  and  supplies  in  the  year  1901  was 
$4510.42  and  the  approximate  cost  of  power  as  noted  above  was 
$1286.83,  making  the  total  operating  expense  $5797.25.  Com- 
puted on  a  ton-mile  basis  the  net  work  costs  were  distributed  as 
follows:  Labor  and  repairs,  1.59c. ;  power  0.46c. ;  total  2.05c. 


HAULAGE  COSTS 


253 


Gross  work  per  ton-mile  costs  were:  Labor  and  repairs,  0.86c. ; 
power,  0.13c. ;  total  1.09c. 


9,500  ft. 
Haul 

10,400  ft. 
Haul 

Total 

Cars  of  coal  

48,120 

14,681 

62801 

Cars  of  rock  

4,233 

235 

4,468 

108,270 

33,032 

141,302 

12,699 

675 

13374 

Total  tons  

120,969 

33,707 

154676 

217,744 

66,600 

284344 

Ton-mileage,  dead  load,  cars  and  empty 

188,471 

58,769 

247,240 

Gross  ton-mileage 

406,215 

125  369 

531  584 

An  excellent  method  of  checking  up  the  work  of  haulage 
motors,  by  means  of  charts  and  graphs,  was  described  at  the 
February,  1915  meeting  of  the  A.  I.  M.  M.  E.  a  typical  chart 
used  by  one  of  the  larger  companies  in  West  Virginia  being 
shown  in  Fig.  10.  As  will  be  noted  the  chart  shows  an  adopted 
standard  of  100  per  cent  efficiency  for  men  and  equipment,  with 
the  curves  showing  what  is  actually  being  accomplished  plotted 
on.  This  particular  chart  shows  the  work  being  done  by  a 
gathering  motor. 

Motor  Losses. — In  any  direct-current  motor,  the  losses  can  be 
divided  under  three  heads.  The  C2R,  or  copper  loss,  is  due  to 
the  resistance  of  the  conductors  on  the  armature  and  field  wind- 
ings. This  loss  varies  with  the  square  of  the  current,  as  will 
be  seen  from  the  formula  C2R  =  watts  lost. 

The  copper  loss  is  composed  principally  of  two  parts,  one 
due  to  the  load  or  line  current  and  the  other  due  to  the  idle 
current  that  circulates  in  the  short-circuited  turns,  of  the 
armature  winding  at  the  instant  of  commutation,  due  to  the 
distortion  of  the  field  by  the  armature  reaction. 

The  core  loss  can  be  separated  into  two  parts.  One  is 
known  as  the  eddy-current  loss  and  is  due  to  the  current  set 
up  in  the  armature  laminations  by  their  being  revolved  in  a 
magnetic  field.  This  loss  increases  with  the  square  of  the  num.- 


254 


COAL  MINING  COSTS 


ber  of  magnetic  lines  per  square  inch,  or  the  density  of  the 
field,  and  with  the  square  of  the  speed. 

The  other  part,  known  as  the  hysteresis  loss,  is  due  to  the  re- 
versals of  the  magnetic  lines  of  force  in    the  armature  iron,  as 


.a 


I 
s 

CJ 

£ 


it  is  being  revolved  in  the  magnetic  field,  which  varies  with  the 
1.6  power  of  the  density  and  directly  with  the  speed.  These 
two  losses  combined  are  known  as  the  core  loss. 

The  third  loss  is  the  mechanical,  or  friction  loss,  and  is  due 
to  the  bearing  friction,  the  brush  friction  on  the  commutator 


HAULAGE  COSTS 


255 


and  the  windage.  This  friction  loss  amounts  to  about  1  per  cent 
at  full  load  of  the  energy  being  supplied  to  the  motor;  and,  as 
it  is  practically  equal  on  all  types  of  machines,  it  can  be  elim- 
inated from  the  discussion. 

The  combined  core  and  copper  losses  in  a  well  designed  motor 
will  amount  to  about  12  to  14  per  cent  of  the  input  at  full  load, 
and  about  8  to  10  per  cent  of  the  input  at  one-half  load.  At  full 
load,  this  loss  is  largely  composed  of  the  C2R  or  copper  loss,  but 
at  one-half  load  it  is  principally  the  core  loss. 

From  this  it  will  be  seen  that  the  copper  loss  is  the  greatest 
loss  at  full  load,  and  the  one  largely  responsible  for  the  one-hour 
rating  of  the  motor.  On  the  other  hand,  it  will  be  noted  that 
for  loads  of  one-half  or  less,  the  core  loss  is  the  principal  loss  and, 
therefore,  largely  responsible  for  the  continuous  rating  of  the 
motor. 

The  copper  loss  of  the  motor  is  well  within  the  power  of  the 
engineer  to  control.  He,  therefore,  can  govern  fairly  well  the 
hourly  rating  that  the  motor  is  to  have.  The  core  loss  of  a 
noncommutating-pole  motor  is  not  so  well  within  the  control  of 
the  engineer,  because  he  must  sacrifice  low  core  loss  in  order  to 
get  good  commutation,  and  in  so  doing,  sacrifice  the  continuous 
rating  of  the  machine. 

The  commutating  pole  minimizes  the  distortion  of  the  field  by 
the  armature  reaction,  fixes  the  brush  position  for  all  loads  with 
either  direction  of  rotation,  prevents  all  sparking  and  burning 
at  the  brushes,  and  reduces  local  current  in  the  armature  winding 
to  a  minimum. 

In  order  to  prove  these  theories  and  substantiate  the  above 
statements,  the  horsepower  time-curves  and  core  loss-curves  of 
two  motors  of  practically  the  same  one-hour  rating  are  shown 
in  Figs.  11  and  12.  These  curves  are  the  result  of  actual  tests, 
conducted  on  these  two  types  of  motors.  The  mechanical  features 
of  the  two  machines  are  as  follows : 


Noncommutating 
Pole  Motor 

Commutating 
Pole  Motor 

Weight  

2130  Ib 

1850  Ib 

Speed  

375  r  p  m 

395  r  p  m 

Mount  on  

28-in.  wheels 

28-in  wheels 

Mount  on  

30-in   firafire 

28-in    era  ffp 

Builder's  rating  

40  hp. 

37£  hp. 

256 


COAL  MINING  COSTS 


Figs.  11  and  12  show  the  horsepower  time  and  core-loss  curves 
of  the  two  motors  on  the  standard  basis  of  the  75-deg.  C.  rise. 

Curve  A  represents  the  noncommutating-pole  motor,  and 
curve  B  the  commutating-pole  machine.  By  referring  to  curve 
Ay  it  will  be  seen  that  the  noncommutating-pole  motor  has  an 
hourly  rating  of  41  hp.,  and  a  continuous  rating  of  16%  hp., 
or  42  per  cent  of  its  one-hour  rating  for  a  continuous  rating. 

Referring  to  curve  B,  it  will  be  seen  that  the  commutating- 
pole  motor  has  a  one  hour  rating  of  38  hp.  and  a  continuous 
rating  of  20.8  hp.,  or  54.8  per  cent  of  its  one-hour  rating  for  a 
continuous  rating. 


3  4 

Time  in  Hours 

IG.  11. — Horsepower  curves  for  the  two  types  of  motors  having 
the  same  1  hr.  rating. 

If  the  noncommutating-pole  motor  had  been  designed  with  in- 
terpoles,  along  the  lines  of  the  commutating-pole  machine,  it 
would  have  54.8  per  cent  of  its  one-hour  rating,  22.6  hp.  for  its 
continuous  rating  instead  of  its  present  rating  of  16!/2  hp.,  an 
increase  of  6.1  hp.,  or  36.4  per  cent  in  continuous  rating. 

Power  Costs. — The  cost  of  power  to  operate  motors  can  be 
reduced  by  obtaining  a  high  load  factor.  The  maximum  energy 
demand  of  the  motors  on  the  grade  will  be  the  same,  since  wo 
presuppose  that  the  largest  number  of  cars  will  be  pulled  at 
all  times,  depending  on  the  grade,  but  the  average  power  con- 
sumption of  the  motor  will  be  increased  at  all  other  points  on 
the  trip,  thus  affecting  the  load  factor  for  the  entire  mine. 


HAULAGE  COSTS 


257 


The  results  of  many  observations  show  that  a  definite  re- 
lation exists  between  the  cost  of  production  of  electrical  energy 
and  the  load  factor.  With  a  small  plant  of  one  unit,  as  usually 
installed  at  a  coal  mine,  the  cost  of  power  at  no  load  is  excessive, 
being  approximately  one-third  to  one-half  that  of  full  load.  On 
account  of  the  terms  of  a  contract  with  the  miners'  union  the 
labor  cost  is  usually  a  constant,  and  the  only  difference  in  the 


CORELOSS\AT    \ 
CONTINUOUS  RAT/ 


CONTINUOUS  RATING. ONE  HOUR  RATING 


ZO    40     60    80   100  120    140  160    180 
Load,  Amperes 

FIG.  12. — Core-loss  curves  of  the  two  types  of  motors  having  the 
same  1  hr.  rating. 


cost  between  no  load  or  part  load  and  full  load  is.  in  the  amount 
of  coal  and  supplies  used.  The  advantage  of  keeping  the  plant 
load  uniformly  near  the  rated  capacity  without  high  peaks  is 
plainly  evident,  and  in  case  the  power  is  purchased  from  a 
central  station,  the  saving  may  be  considerable  by  obtaining  a 
sliding  scale  contract  providing  for  this  feature. 

In  order  to  express  the  total  value  of  the  power  saved  by  a 
formula : 


258  COAL  MINING  COSTS 

Let          n  =  the  number  of  trips  made  by  the  locomotive  per 

shift; 
g=  gross  weight  of  locomotive  and  trips,  both  loaded 

and  empty,  when  the  mine  is  producing  t  tons 

of  coal; 
G  =  gross  weight  of  locomotive  and  trips,  if  mine  pro- 

duced T  tons  of  coal  through  adding  x  cars  to 

each  trip; 
(T—t)  =  (G—g)n  =  increased  tonnage  hauled; 

L  =  net  tonnage  of  locomotive  alone  for  all  trips; 
ET  =  the  efficiency  or  percentage  of  power  used  in  haul- 

ing cars  when  the  mine  produces  T  tons: 
Et  =  ihe  efficiency  when  the  mine  produces  t  tons; 
K  =  cost    of   power   per   kilowatt-hour   expressed   in 

dollars; 
R  =  per  cent  saved  per  kilowatt-hour  due  to  reduction 

of  load  factor; 

PT  =  power  used  to  produce  T  tons; 
Pt  =  power  used  to  produce  t  tons; 
Si  =  saving  from  increased  efficiency  of  power  resulting 

from  a  reduced  ton-mileage  of  the  locomotive; 
82  =  saving  per  ton  from  reduction  of  the  load  factor; 
$3  =  saving  per  ton  due  to  lowered  summits  or  of  rise 

and  fall,  acceleration  and  braking  power  losses 

not  being  considered; 
Sp  =  total  saving  per  ton  of  coal  produced,  in  power 

cost  resulting  from  an  increased  tonnage  pro- 

duced by  reducing  the  limiting  grade. 

Then: 

ET=(nG—L)/Gn  when  producing  T  tons; 
Et=(ng—L)/gn  when  producing  t  tons; 

51  =  (Et-ET)KPT/T; 

52  =  RKPT/T; 


Sp  =  [(Et-ET)+R]KPT/T-(Pt/t-PT/T)K. 

The  following  is  a  typical  example  of  load  allowances  on  a 
motor  assuming  a  4000  ft.  haul  with  a  stiff  grade  against  the 
loads  the  full  length  :  The  haul  up  this  grade  at  normal  speed 
will  require  10  to  12  min.,  under  about  three-fourths  full  power; 


HAULAGE  COSTS  259 

switching,  coupling  to  empties,  etc.,  will  require  1  to  2  min. ; 
the  return  trip  with  the  empties  will  take  10  min.  with  an  ad- 
ditional 1  or  2  min.  to  couple  up  to  the  load.  The  round  trip  thus 
requires  from  22  to  26  min.  During  this  interval  the  motor  will 
be  under  three-quarter  load  for  12  min.,  little  or  no  load  for  3  or 
4  min.,  and  10  min.  on  the  return  trip  under  about  one-quarter 
load. 

At  the  mines  of  the  Copper  Queen  Mining  Co.,  in  Bisbee,  the 
power  used  on  trolley  locomotives,  measured  at  direct-current 
switchboard  in  power  station,  for  the  year  1912  amounted  to  875 
watt-hours  per  useful  ton-mile.  This  amount,  however,  includes 
a  few  lights  which  are  connected  to  the  trolley  circuit  and  gives 
too  high  a  figure  for  the  locomotives  alone.  It  applies  to  cars 
with  roller  bearings,  about  one-half  of  the  tonnage  being  carried 
in  cars  of  two  tons  capacity  and  the  other  half  in  cars  of  one 
ton  capacity.  The  conditions  of  the  cars  and  tracks  have  quite 
an  important  bearing  on  power  required  per  ton-mile. 

For  the  year  1912  the  cost  of  various  items  in  cents  per  useful 
ton-mile  at  Bisbee  for  a  total  of  408,000  ton-miles  was  as  follows : 


Cents 

Locomotive  maintenance 2 . 95 

Car  maintenance 1 . 64 

Track  maintenance 5 . 24 

Trolley  maintenance 3 . 60 

Power..  .   1.64 


Locomotive  maintenance  includes  all  electrical  and  mechan- 
ical repairs  and  replacements  on  locomotives,  as  well  as  lubri- 
cating oil  and  supplies. 

Car  maintenance  includes  all  repairs,  oil  and  supplies  on  cars. 

Track  maintenance  includes  all  track  repairs  and  replace- 
ments, bonding,  grading  and  realignment. 

Trolley  maintenance  includes  all  trolley-wire  repairs  and  re- 
placements, and  repairs  to  protective  trough  around  trolley  wire. 

Track  and  trolley  maintenance  are  very  heavy,  due  to  shifting 
ground. 

The  cost  of  power  (1.1  kw-hr.  per  ton-mile)  is  taken  at  the 
high-tension  switchboard  and  includes  the  loss  in  transforming 
the  alternating  current  into  direct  current. 


260 


COAL  MINING  COSTS 


In  comparing  power  used  by  storage-battery  locomotives  and 
trolley  locomotives  it  would  be  fairer  either  to  compare  the  actual 
input  into  battery  with  direct-current  power  used  by  trolley 
locomotives  or  use  the  alternating-current  power  input  to  rotary 
converter  in  both  cases. 

In  the  case  of  the  storage  battery  the  input  was  approximately 
1.28  kw-hr.  per  ton-mile,  as  against  0.875  kw-hr.  for  the  trolley 
locomotive. 

The  figures  for  power  on  storage  battery  are  based  on  two 
days '  test  and  therefore  are  not  as  reliable  as  those  on  the  trolley 
locomotives,  which  cover  a  year's  period. 

The  accompanying  tables,  the  figures  of  which  are  as  of  1917, 
may  be  used  as  the  basis  of  power  computation  costs  in  connection 
with  the  computation  of  haulage  costs.  They  are  sufficiently  close 
to  give  the  approximate  interest  and  depreciation  charges  on 
a  deferred  expenditure  for  equipment,  made  possible  by  reducing 
grades  and  thus  lowering  the  maximum  energy  demands  at 
peak  loads,  and  for  other  purposes  of  an  approximate  nature : 


NUMBER  OP  KILOWATTS  REQUIRED  TO  HAUL  ONE  TON  GROSS  TRAIN  LOAD 
(As  MEASURED  AT  THE  TROLLEY  WIRE) 

The  total  resistance  includes  the  tractive  resistance  of  20  and  30  Ib.  per 
ton  on  level  track  and  grade  resistances  up  to  4  per  cent,  inclusive. 

Velocity  in  Miles  per  Hour 


4 

5 

6 

"       8 

10 

Tractive 

Tractive 

Tractive 

Tractive 

Tractive 

Per  Cent 

Resistance, 

Resistance, 

Resistance, 

Resistance, 

Resistance, 

Grade 

Pounds 

Pounds 

Pounds 

Pounds 

Pounds 

per  Ton. 

per  Ton. 

per  Ton. 

per  Ton 

per  Ton. 

20       30 

20        30 

20        30 

20       30 

20         30 

0 

0.2     0.3 

0.25     0.38 

0.3      0.45 

0.4     0.6 

0.50     0.75 

0.5 

0.3     0.4 

0.38     0.50 

0.45     0.60 

0.6     0.8 

0.75     1.00 

1.0 

0.4     0.5 

0.50    0.63 

0.60    0.75 

0.8     1.0 

1.00     1.25 

1.5 

0.5     0.6 

0.63     0.75 

0.75    0.90 

1.0     1.2 

1.25     1.50 

2.0 

0.6     0.7 

0.75    0.88 

0.90       .05 

1.2     1.4 

1.50     1.75 

2.5 

0.7     0.8 

0.88     1.00 

1.05       .20 

1.4     1.6 

1.75     2.00 

3.0 

0.8     0.9 

1.00     1.13 

1.20       .35 

1.6     1.8 

2.00    2.25 

3.5 

0.9     1.0 

1.13     1.25 

1.35       .50 

1.8     2.0 

2.25     2.50 

4.0 

1.0     1.1 

1.25     1.38 

1.50       .65 

2.0    2.2 

2.5       2.75 

HAULAGE  COSTS 


261 


Formula  : 


Kw  =  kilowatts  used  to  haul  one  ton  as  measured  at  the  trolley 
wire.     (The  losses  in  the  line  must  be  added  to  obtain 
the  total  power  indicated  at  the  switchboard.) 
Tr  =  tractive  resistance  in  pounds  per  ton; 
v  =  velocity  in  feet  per  second. 

NOTE.  —  The  efficiency  of  the  motor  and  gearing  are  taken  as  80  per  cent 
when  running  with  the  controller  cut  out.  With  the  controller  in  circuit,  the 
efficiency  of  the  motor  will  be  from  60  to  65  per  cent. 

AVERAGE  COST  PER  KILOWATT  PRODUCED  FOR  THE  VARIOUS  POWER-PLANT 

UNITS 


Item 

Minimum  Cost 
per  Kilowatt 
Capacity    . 

Maximum  Cost 
per  Kilowatt 

Boilers  and  settings 

$10  75 

$13  25 

Stokers 

1  30 

2  20 

Flues,  dampers  and  regulators  
Boiler-feed  pumps  

.60 

.40 

.90 
.75 

Feed-water  heater.  . 

.20 

35 

Piping,  valves,  etc 

4  20 

7  00 

Economizers  
Foundations  for  engines 

1.30 
2  00 

2.25 
3  00 

Engines  

22.00 

30.00 

Generators  
Switchboards  and  wiring 

16.60 
3  20 

22.80 
4  20 

Supervision  

4.00 

6  00 

Miscellaneous 

2  00 

3  00 

Total  

$68  55 

$95  70 

A  typical  round-trip  performance  of  a  mine  locomotive  work- 
ing actual  mine  conditions,  is  indicated  by  curves  shown  in 
Fig.  13.  This  is  a  6-ton  locomotive  with  two  tandem  motors, 
rated  at  6  miles  per  hour,  with  an  electrical  input  of  40  kw.  on 
a  drawbar  pull  of  2400  Ib.  or  400  Ib.  per  ton. 

The  starting  drawbar  pull,  or  maximum  tractive  effort,  ie> 
approximately  150  per  cent  of  that  of  3600  Ib.,  with  a  current, 
consumption  of  500  v.  of  115  amp.  The  ragged  current  and  volt- 
age curves  are  noted  to  occur  during  the  whole  of  the  round-trip. 


262 


COAL  MINING  COSTS 


The  current  serves  also  as  a  direct  measure  of  the  torque,  which 
can  be  computed  from  the  actual  characteristic  curves. 

As  the  entry  in  this  mine  often  runs  on  the  strike  of  the  coal, 
one  side  of  the  track  is  ballasted  and  the  other  rests  on  bed 
rock.  This  means  a  poor,  uneven  track,  sometimes  dirty  from 
loose  coal  falling  off  cars  or  from  rock  falls.  Then,  too,  wet  spots 
and  small  increases  of  grades  are  often  met  with,  requiring  fre- 
quent sanding.  Although  the  couplings  are  long,  the  grade 


I      of 

6  TOWtf'JVe  LOCOHOT/VC.. 
?-  SOO  V:  /V-  Motors.    4O-t(^W  °  \ 
*5  °  C.  ftfse  -  43/tfnp.  for.  /  Hour. 

\&f?-6M.PH. 
Whee/  O/0/n.  28  /nchts. 
Gear  ffaf/'o  /S/o  94. 
L/'fte  /fes.  2  Oftma. 


60   .70     SO    9O    JOO 


A/np.  O     10    ZO    30 

FIG.  13.  —  Typical  round-trip  performance  of  a  mine  locomotive  working 
under  actual  mine  conditions. 

keeps  them  tight  and  requires  a  greater  accelerative  torque  than 
when  running  on  the  level. 

As  this  entry  is  one  of  four  in  a  mine  about  2  miles  from  the 
power  station,  and  each  of  them  has  a  locomotive,  besides  using 
four  electric  booster  fans  and  an  electric  hoist,  no  further  ex- 
planation is  necessary  of  the  apparently  inconsistent  variation 
of  voltage.  It  is  typical  of  mine  service.  The  generators  at  the 
power  house  are  not  overcompounded,  and  there  is,  therefore, 
no  compensation  for  line  loss.  This  results  in  this  particular  case 
in  quite  a  drop  of  pressure  when  the  electric  locomotive  is  pulling 


HAULAGE  COSTS  263 

heavily ;  consequently,  the  characteristics  obtained  at  500  v.  can 
not  be  very  well  used  for  correct  calculation  at  100  amp.  con- 
sumption. The  changes  in  such  characteristics  due  to  two  ohms 
drop  in  the  feeder  line  and  return  are  sketched  in  roughly.  The 
decrease  in  electrical  and  brake  horsepower  is  very  apparent, 
resulting  in  great  loss  of  speed  at  heavy  loads,  but  increasing  the 
tractive  effort  at  this  lower  speed. 

It  is  interesting  to  note  the  enormous  difference  of  perform- 
ance between  an  ordinary  day  trip  such  as  is  shown  here,  and 
one  with  very  cold  weather  outside.  The  empties  coming  from 
the  outside  run  with  great  difficulty  because  of  frozen  boxings, 
and  often  require  more  than  a  50-per  cent  increase  in  torque; 
by  the  time  they  are  loaded,  however,  the  mine  temperature  has 
thawed  them  sufficiently  to  make  the  outgoing  trip  a  fairly 
normal  one. 

Where  the  inertia  of  the  trip  of  cars  at  starting,  or  increased 
resistance  due  to  a  dirty  track,  causes  the  locomotive  to  pull 
heavily  and  run  slowly,  the  effect  of  a  long  feed  line  causing 
an  appreciable  voltage  drop  at  the  locomotive  is  of  interest.  The 
drop,  of  course,  is  proportional  to  the  current  consumption.  The 
motor  torque  is  proportional  to  the  current  only,  and  is  inde- 
pendent of  the  impressed  electromotive  force.  With  a  heavy 
pull,  the  motor 's  speed  need  not  be  so  great;  as  an  increase  in 
current  flow  means  less  speed  for  the  necessary  current  electro- 
motive force.  As  the  applied  electromotive  force  decreases  (due 
to  line  loss)  and  the  current  electromotive  force  increases  with 
the  greater  current  flow,  the  torque,  and  therefore  the  current, 
does  not  increase  as  in  the  case  of  constant  applied  voltage. 
There  is,  therefore,  less  danger  of  overloading  the  motors.  As 
the  tractive  effort  also  is  higher  at  lower  speeds,  such  line  resist- 
ance allows  of  larger  trips  to  be  started  by  a  certain  weight  loco- 
motive without  an  excessive  output  at  the  power  house.  As  the 
trip  resistance  is  decreased  after  starting,  the  speed  rises  with  the 
decreased  current  and  increased  applied  electromotive  force 
making  up  for  time  lost  on  the  heavy  pull  when  starting. 

Another  advantage  of  a  long-feed  line  is  decreased  harmful- 
ness  of  short-circuits.  A  short-circuit  on  this  2-mile  line  will  blow 
the  circuit  breaker  without  causing  a  violent  flashing  of  the  gen- 
erator brushes,  or  excessive  arching  at  the  contacts  of  the  circuit 
breaker,  since  the  current  cannot  exceed  300  amp. 


264  COAL  MINING  COSTS 

Line  costs  and  losses. — Line  losses  can  best  be  shown  by  a 
practical  example.  For  purposes  of  demonstration  it  will  be 
assumed  that  the  haulage  road  at  a  certain  mine  is  6000  ft. 
long  and  laid  with  30-lb.  rail,  which  are  bonded  with  00  wire, 
and  bonding  caps  with  cross  bonds  every  30  ft.  There  is  a  sub- 
station at  the  inside  end,  250  volt  current  is  used,  the  trolley 
wire  is  0000  and  a  10-ton  motor  using  current  at  150  amp.,  is 
used.  Let  it  be  desired  to  find: 

What  will  be  the  drop  in  volts  at  2000  ft.,  4000  ft.,  and  6000 
ft.  from  the  substation. 

What  is  the  percentage  of  loss  in  each  case. 

What  would  be  the  saving,  if  any,  in  installing  0000  feeder 
wire,  allowing  say  6  per  cent  interest  on  the  investment  and  com- 
puting the  cost  of  power  at  2c.  per  kw.-hr. 

In  estimating  the  drop  at  distances  of  2000,  4000  and  6000 
ft.  from  the  substation,  it  is  necessary  to  calculate  the  trolley- 
drop  and  rail-drop  separately  and  add  the  results.  Then,  in 
order  to  estimate  the  possible  advantage  of  installing  a  feeder-line 
of  the  same  size  as  the  trolley-wire  (0000),  it  is  necessary  to  cal- 
culate further  the  combined  trolley  and  feeder-drop  for  the 
same  distances. 

The  first  step  in  the  calculation  is  to  take  from  an  electric- 
wire  table  the  circular  mils  of  a  0000  copper  wire,  which  is 
211,600  circ.mils.  The  combined  trolley  and  feeder  wires,  the 
circular  mils  is  double  this  amount,  or  423,200  circ.mils.  Now, 
to  find  the  copper  equivalent  of  two  30-lb.  rails,  properly  bonded 
and  cross-bonded,  assume  a  rail-to-copper  ratio  of  11 :1 ;  that  is  to 
say,  take  the  rail  resistance  here  as  11  times  that  of  copper,  for 
the  same  sectional  area.  Then  find  the  combined  sectional  area 
of  two  30-lb.  rails  as  follows:  Since  the  weight  of  wrought 
iron  is  480  Ib.  per  cu.ft,  and  two  30-lb.  rails  weigh  60  Ib.  per 
lineal  yard  (36  in.),  the  cubic  contents  of  the  two  rails  is 
60/480X1728=216  cu.in.  per  yd.  The  corresponding  sectional 
area  for  the  two  rails  is  therefore  216-^-36=6  sq.in. 

Again,  since  1  mil  is  1/1000  in.,  and  a  circ.  mil  is  the  area  of  a 
circle  whose  diameter  is  1  mil,  1  circ.mil  =  0.7854  (1/1000)2  = 
0.0000007854  sq.in.  and  1  sq.in.  =  1/0.0000007854  =  1,273,200 
circ.  mils.  The  sectional  area  of  the  two  30-lb.  rails  in  circular  mils 
is  therefore  6X1,273,200  =  7,639,200  circ.mils.  For  a  rail-to- 


HAULAGE  COSTS  265 

copper  ratio  of  11  :  1,  the  copper  equivalent  is  7,639,200  -^  11  = 
say,  694,500  circ.mils  of  copper. 

It  is  now  possible  to  calculate  the  drop  of  potential  for  the 
trolley  line,  feeder  and  rail  return  for  each  required  distance  by 
using  the  general  formula: 

~          .          ,.  ,     10.8  (distanceX  current) 
Drop  of  potential  =  --  -  -  :  --  =  -  -. 

ctrc.mils 

The  drop  of  potential  is  expressed  in  volts  and  the  current  in 
amperes,  while  the  distance  is  given  in  feet.  Applying  this  formula 
to  the  present  case  gives  for  the  drop  at  a  distance  of  2000  ft. 
from  the  substation  the  following: 

10.8(2000X150) 
Trolley-drop  =  --  --  =  15-3  voUs- 


It  is  evident  that  the  combined  trolley-  and  feeder-drop  would 
be  one-half  of  this  amount,  or  7.65  volts,  since  there  are  two  wires 
of  equal  size  for  the  transmission  of  the  current  to  the  locomotive. 
The  rail-drop  is: 

D   „  ,          10.8(2000X150) 
Rail-drop  =  -  ^=4.6  vote. 


The  above  results  give  the  drop  of  potential  for  a  distance  of 
2000  ft.      For  distances  of  4000  and  6000   ft.  the  drop  will  be 
two  and  three  times  the  above  amount  respectively. 
A  comparison  of  these  results  gives  the  following  : 

Total  drop  without  feeder,      4.6+15.3   =      19.9    volts 
Total  drop  with  feeder,  4.6+  7.65=      12.25  volts 

Voltage  saved  by  feeder,       19.9  -  12.25  =        7.65  volts 
Wattage  saved  by  feeder,        150X7.65  =  1147.5    watts 
The  percentage  of  drop  in  each  case  above  is: 

Without  feeder  (2000  ft.),  19.9X100-^250  =  8.0% 

With  feeder  (2000  ft,),  12.25X100^-250=4.9% 

Assuming  250  working-days  of  8  hr.  each  (2000  working  hours) 

in  a  year,  the  saving  in  that  time  at  a  cost  for  electricity  of  2c. 

per  kw.-hr.  would  be  2000X1.  1475X0.02  =  $45.90. 

Against  this  saving  must  be  reckoned  the  interest  or  fixed 
charges  on  the  investment  for  the  erection  of  2000  ft.  of  feeder 
wire.  The  weight  of  0000  copper  wire  is  practically  640  Ib. 
per  1000  ft.  making  the  costs  for  2000  ft.  of  this  wire,  at  20c. 


266  COAL  MINING  COSTS 

per  lb.,  2  X  640  X  0.20  =  $2.56.  The  cost  of  erection,  including 
poles,  will  be  about  $4  per  100  ft.,  or  $80  for  a  distance  of 
2000  ft.,  which  makes  the  total  cost  of  this  length  of  feeder 
wire  erected,  $336. 

Estimating  the  fixed  charges  on  this  investment  as,  interest, 
6  per  cent;  depreciation,  4  per  cent,  and  taxes  and  insurance, 
2  per  cent,  making  a  total  of  12  per  cent,  gives  0.12  X  336  = 
$40.32.  This  estimate  makes  the  net  saving  per  year,  $45.90— 
40.32=$5.58. 

Since  the  drop  in  potential  the  saving  in  kilowatt-hours  and 
the  fixed  charges  are  all  proportional  to  the  distance,  the  net 
saving  per  year  will  also  be  proportional  to  the  distance,  or 
$11.16  for  a  distance  of  4000  ft.,  and  $16.74  for  a  distance  of 
6000  ft. 

In  actual  practice,  however,  the  locomotive  .will  not  be  run- 
ning more  than  from  one-half  to  three-quarters  of  the  actual 
time,  depending  on  the  length  of  the  haul  and  other  conditions, 
which  will  reduce  the  saving  in  kilowatt-hours  at  the  same  rate, 
while  the  fixed  charges  must  remain  constant.  Although  the 
figures  would  then  show  an  excess  of  fixed  charges  over  the 
saving  in  the  cost  of  electricity,  it  would  still  be  good  practice 
to  add  the  feeder  wire.  The  gain  in  the  operation  of  the  loco- 
motive will  exceed  what  the  actual  saving  in  power  would 
indicate. 

Grounding  losses  can  sometimes  be  discovered  on  the  ammeter 
when  the  mine  is  shut  down  or  at  any  other  time  when  all  the 
power  is  supposed  to  be  off.  In  one  instance  an  investigation 
of  a  75  amp.  reading  of  the  ammeter  when  the  mine  was  supposed 
to  be  closed  for  the  day  disclosed  the  fact  that  25  trolley  wire 
hangers  were  grounded.  When  the  trouble  was  remedied  the 
ammeter  read  zero. 

A  simple  calculation  will  show  the  loss  from  this  leakage  of 
75  amp.  under  a  pressure  of  275  volts,  being  .continuous  for  the 
time  the  switch  on  the  supply  current  is  open,  which  is  16  hr. 
a  day.  This  loss  in  six  months  of  125  working-days  of  16  hr. 
each,  or  say  2000  hr.,  would  amount  to  (75  X  275  X  2000) 
-4- 1000  =  41,250  kw.-hr.  At  a  cost  of  2c.  per  kw.-hr.,  the  loss 
would  be  41,250X0.02=$825  in  six  months.  In  the  present  case, 
the  loss  being  due  to  the  leakage  of  current  by  the  grounding 
of  25  trolley- wire  hangers,  the  loss  is  825  -f-  25  =  $33  per  hanger 


HAULAGE  COSTS  267 

in  six  months,  or  $5.50  per  hanger  per  month — an  amount  that 
is  almost  inconceivable.  However,  it  shows  the  importance 
of  carefully  testing  an  electric  circuit  for  grounds  and  locating 
and  stopping  the  leak.  Besides  this  direct  loss  of  current,  there 
is  the  danger  from  fire  where  the  hangers  are  in  the  coal  roof 
or  insulator  pins  are  used  instead  in  the  entry  ribs. 

The  following  is  an  approximate  estimate  of  the  cost  of  1000 
ft.  of  trolley  construction  in  coal  mines,  including  rail  bonds  as 
estimated  in  1908: 


36  single-bolt  roof  suspensions $15 

30  straight  ears  for  grooved  wire . 6 

6  curved  ears  for  grooved  wire 6 

1000  ft.  of  0000  grooved  trolley  wire  (641  Ib.) 94 

210  ft.  of  00  bond  wire  (85  Ib.) 12 

175  ft.  channel  pins,  GE  Catalog,  No.  17,315 5 

Labor  and  other  material 50 

Additional  labor,  if  necessary,  to  drill  rails  for  bonds  with  a  hand 

drill.  .  15 


Total  cost  per  1000  ft $197 

Other  sizes  of  trolley  and  bond  wires  may  be  readily  sub- 
stituted in  the  above  estimate. 

Bonding. — While  a  few  operators  are  using  modern  methods 
of  bonding,  many  are  depending  on  channel  pins  and  wire 
to  carry  the  return  circuit.  The  channel  pin  when  first  installed 
is  more  or  less  efficient,  but  as  three-fourths  of  its  contact  is 
between  steel  pin  and  steel  rail,  it  is  impossible  to  obtain  a  union 
that  will  exclude*  air  and  moisture.  Therefore  it  is  only  a  short 
time  until  corrosion  has  started  and  a  high  resistance  is  intro- 
duced at  the  points  .of  contact.  The  method  of  testing  prevalent 
at  most  mines  consists  of  examining  the  bond  to  see  that  the 
wire  and  pins  are  intact. 

The  return  circuit  of  a  mine  that  was  bonded  partly  with 
channel  pins  and  partly  with  compressed-terminal  flexible-cable 
bonds  was  tested  with  a  direct-reading  bond  tester  which  showed 
the  resistance  of  each  joint  as  equal  to  the  resistance  of  a  certain 
length  of  solid  rail.  Thirty-one  per  cent  of  the  channel-pin 
bonds  showed  a  resistance  equal  to,  or  greater,  than  that  of  a 
30  ft.  rail,  or  practically  an  open  joint;  the  average  resistance 
of  the  balance  was  equal  to  that  of  13  ft.  of  rail.  These  channel- 


268  COAL  MINING  COSTS 

pins  had  been  installed  about  2%  yr.  and  on  an  exceptionally 
good  roadbed. 

The  compressed  terminal  bonds  had  been  installed  4  yr.  on  a 
roadbed  that  was  in  bad  condition.  The  drainage  was  bad  and 
the  soft  roadbed  permitted  a  considerable  rising  and  sinking  of 
each  joint  when  a  car  passed  over,  thus  imposing  a  severe  strain 
on  the  bond  terminals;  16  per  cent,  of  these  bonds  were  found 
defective,  and  the  balance  showed  an  average  resistance  of  6.6  ft. 
Had  these  bonds  been  installed  in  track  similar  to  that  in  which 
the  channel-pins  were  used  there  is  no  doubt  but  that  the  depre- 
ciation would  have  been  cut  in  half. 

A  point  of  special  interest  in  this  test  was  the  fact  that  the 
majority  of  the  compressed-terminal  bonds  were  in  good  condi- 
tion after  4  yr.  of  service  and  under  rather  unusually  un- 
favorable conditions.  Had  this  company  tested  these  bonds  at 
certain  intervals,  and  replaced  defective  ones  as  found,  they 
would  have  had,  at  a  small  expense  for  labor  and  material,  a 
highly  efficient  return  circuit  at  all  times. 

This  operation  had  been  suffering  from  an  excessive  drop 
in  voltage,  and  the  results  of  this  test  proved  conclusively  that 
it  was  caused  by  a  defective  transmission  of  the  return  current. 

The  relative  resistances  of  21  joints  selected  at  random  along 
the  haulage  system  of  this  mine,  bonded  first  with  channel- 
pin  bonds  then  with  compressed-terminal  bonds,  are  clearly 
shown  in  the  accompanying  curve,  Fig.  14. 

One  mine  manager  reported  that,  with  sufficient  copper  over- 
head, he  found  a  drop  of  100  volts,  at  less  than  3000  ft.  from 
the  generator  on  250-volt  direct-current  circuit.  The  channel- 
pin  bonds  in  this  mine  were  tested,  and  90  per  cent  of  them 
found  to  have  a  resistance  greater  than  that  of  30  ft.  of  rail. 
The  channel-pins  were  replaced  by  compressed-terminal  bonds, 
the  line  voltage  went  up  to  normal  and  the  efficiency  of  the 
locomotive  increased  to  a  marked  extent. 

Computing  losses  in  bonds.— It  is  essential,  both  in  making 
tests  of  bond  installations,  and  in  estimating  new  work  that  the 
resistance  of  a  bonded  joint  be  known.  The  following  tables  and 
formula  show  just  what  resistance  a  well  bonded  joint  will  have. 

To  find  the  resistance  of  a  rail  joint  bonded  with  compressed 
terminal  bonds,  the  following  formula  is  used: 
LXR+2XCR 

~RF~      =JR 


HAULAGE  COSTS 


269 


in  which       L  = length  of  bond  in  inches; 

R  =  resistance  of  one  inch  of  cable  or  strands  compos- 
ing the  bond; 

CR  =  Contact  resistance  of  bond  terminals; 
RF  =  Resistance  of  one  foot  of  rail; 
JR  =  resistance  of  bonded  joint  expressed  in  feet  of  rail. 

As  most  bond-testing  instruments  show  the  resistance  of  the 
bonded  joint,  as  compared  with  the  equal  resistance  of  a  certain 


FIG.  14. — Results  of  21  tests  of  channel-pin  and  compressed-terminal  bonds. 

number  of  feet  of  unbroken  rail,  it  is  more  convenient  to  express 
this  value  in  such  terms  than  in  ohms. 

To  illustrate  the  use  of  the  above  formula  and  tables,  assume 
that  a  40-lb.  T-rail  is  to  be  bonded  with  compressed  terminal 
flexible  cable  bonds,  4/0  capacity,  %-in.  terminals,  26  in.  in 
length.  The  resistance  of  each  joint  when  the  bond  has  been 
installed  is  desired.  By  referring  to  the  tables,  the  resistance 
of  one  inch  of  4/0  cable  is  found  to  be  0.00000414  ohm,  and  the 
resistance  of  a  26-in.  bond  is  26  times  0.00000414,  or  0.00010764. 
The  resistance  of  a  %-in.  terminal  is  0.00000053  ohm,  and  of 
the  two  terminals  is  0.00000106  ohm ;  adding  the  cable  resistance 
and  the  contact  resistance  the  total  ohmic  resistance  of  the  in- 
stalled bond  is  0.0001087  ohm.  To  express  this  in  terms  of 


270 


COAL  MINING  COSTS 


RESISTANCE  AND  CARRYING  CAPACITY  OF  RAILS 

Figures   Based   on    Rails    Having  a  Ratio  of  12  to  1,    as   Compared  with 
Copper,  and  at  70°  F. 


Weight, 
Pound  per  Yard 

Resistance, 
Ohms  per  Foot 

Carrying 
Capacity  in  C.M. 

16 

0.0000622 

169,764 

20 

0.00004923 

212,206 

25 

0.00003935 

265,257 

30 

0.00003321 

318,309 

35 

0.00002844 

371,360 

40 

0.00002489 

424,412 

45 

0.00002212 

477,463 

50 

0.000019355 

530,515 

60 

0.0000166 

636,618 

RESISTANCE  OF  SOLID  TERMINALS 
Figures  Based  on  a  Pressure  of  15  Tons  per  Square  Inch  of  Contact  Surface 


Diameters 

Resistance, 
Ohms 

1" 

0.0000008 

f" 

0.00000064 

*" 

0.00000053 

f" 

0.00000045 

1" 

0.0000004 

RESISTANCE  OF  BOND  CABLES  PER  INCH  OF  CONDUCTOR  AT  75°  F. 


Size 

Resistance, 
Ohms  per  Inch 

Capacity  in 
Amperes 

1/0 

0.00000829 

210 

2/0 

0.00000657 

265 

3/0 

0.00000521 

335 

4/0 

0.00000414 

425 

250,000  C.M. 

0.0000035 

500 

300,000  C.M. 

0.00000275 

600 

350,000  C.M. 

0.0000025 

700 

400,000  C.M. 

0.00000219 

800 

HAULAGE  COSTS  271 

equivalent  rail  length,  divide  by  the  resistance  of  1  foot  of 
40-lb.  rail,  and  the  resistance  of  the  bonded  joint  will  be  found 
to  equal  that  of  4.2  ft.  of  unbroken  40-lb.  rail. 

Trials  made  with  standard  bond-testing  instruments  show  that 
when  bonds  are  installed  with  reasonable  care,  the  resistance 
of  the  joint  will  very  closely  approximate  this  calculated  re- 
sistance. Abutting  rail  ends  and  clean  tight  splice  bars  may 
slightly  lower  the  resistance  of  the  joint,  but  this  is  quite  neg- 
ligible and  should  be  disregarded.  Numerous  tests  made  of  com- 
plete haulage  roads,  installed  under  usual  mining  conditions, 
show  that  an  average  will  vary  but  a  fraction  of  a  foot  from 
the  calculated  resistance. 

The  use  of  the  tables  will  not  only  be  found  of  benefit  in  learn- 
ing standards  to  test  individual  joints  for  their  efficiency,  but 
they  can  be  used  to  great  advantage  in  figuring  voltage  drop 
on  proposed  work.  In  calculating  voltage  drop  on  a  circuit  com- 
posed of  a  trolley  and  a  rail  return,  a  formula  is  generally  used 
which  is  correct  for  the  copper  loss,  but  does  not  take  into 
consideration  the  weight  of  the  rail  and  the  size  and  length 
of  the  bonds.  By  calculating  the  voltage  drop  on  the  trolley 
and  feeders  only,  and  then  on  the  rail,  when  properly  bonded, 
the  exact  drop  for  a  given  load  is  secured.  This  method  has 
been  found  particularly  advantageous  where  a  potential  of 
250  volts  is  used  and  the  current  transmitted  over  long  dis- 
tances, as  is  common  with  bituminous  mines. 

For  instance,  taking  the  example  mentioned  above,  assume 
that  a  road  3000  ft.  long  is  to  be  bonded,  and  the  actual  drop 
on  the  return  side  of  the  circuit  is  desired.  With  30-ft.  rail 
lengths,  there  will  be  100  joints  on  one  rail,  having  a  total 
resistance  of  420  ft.  About  10  per  cent  should  be  added  to 
the  joint  resistance  to  take  care  of  short  rail  lengths  and  bond- 
ing at  switches. 

Then  the  resistance  of  one  rail  will  be  equal  to  that  of 
3460  ft.  of  unbroken  rail,  or  0.00836394  ohm,  or  for  the  two 
rails  in  parallel  0.00418197  ohm.  The  voltage  loss  on  the  return 
side  is  the  load  times  the  resistance,  then  by  calculating  the 
drop  on  the  trolley  side,  the  exact  drop  on  the  circuit  is  easily 
ascertained. 

In  the  same  manner,  these  tables  can  be  used  to  test  the 
efficiency  of  the  bonding  of  an  entire  haulage  road  on  a  certain 


272  COAL  MINING  COSTS 

section.  With  voltmeters  at  both  ends  of  a  section  and  an 
ammeter  on  a  locomotive,  the  voltage  drop  at  a  certain  load 
can  be  obtained.  It  is  a  simple  matter  to  calculate  the  drop 
on  the  positive  side  of  the  circuit  and  on  the  return  side,  assum- 
ing the  bonds  are  in  good  condition.  Any  difference  found 
will  be  an  increased  joint  resistance,  and  the  individual  bonds 
should  be  tested  with  a  bond  tester  and  the  defective  ones 
replaced. 

A  number  of  companies  have  adopted  this  method,  as  an 
entire  mine  can  be  gone  over  on  an  idle  day  or  night,  and  the 
individual  joints  need  only  to  be  tested  in  those  sections  which 
show  defective  bonding.  Mine  operators  who  are  desirous  of 
obtaining  efficient  and  economical  results  from  electrical  min- 
ing equipment  will  find  it  well  worth  their  while  to  make  such 
tests  following  them  up  with  any  necessary  repairs. 

One  company  recently  tested  in  this  manner  a  newly  bonded 
road.  The  test  showed  that  the  joint  resistance  was  69  per 
cent  of  the  total  circuit  resistance,  where  it  should  have  been 
only  8  per  cent.  Upon  making  an  examination  of  the  road, 
it  was  found  that  the  track  men  had  neglected  to  install  bonds 
at  two  of  the  switches. 

As  their  load  was  heavy,  this  unnecessary  resistance  would 
have  cost  a  considerable  sum  in  a  month's  time  for  power, 
which  in  this  instance  was  purchased. 

The  following  is  another  example  of  computing  bonding 
losses  it  being  assumed  that  it  is  desired  to  know  what  is  the 
difference  in  resistance  between  a  4-0  round  copper  wire 
3000  ft.  long  and  the  two  25-lb.  steel  rails,  in  a  track,  2000  ft. 
long,  followed  by  two  30-lb.  steel  rails,  in  1000  ft.  of  track. 
The  rails  are  bonded  with  pressed-terminal  all-wire  2-0  bonds. 
.  Also  determine  how  many  kilowatt-hours  will  be  available 
at  the  end  of  a  250-volt  line,  where  a  30-hp.  motor,  3000  ft. 
from  the  generator,  is  taking  100  amperes. 

The  resistance  of  1000  ft.  of  a  4-0  copper  wire  at,  say 
68  deg.  F.  (20  deg.  C.),  as  taken  from  a  table  giving  the  resist- 
ances of  copper  wires  for  different  gages  and  temperatures,  in 
ohms  per  thousand  feet,  is  0.04893  ohm.  The  resistance  for  such  a 
conductor  3000  ft.  long  is  then,  3  X  0.04893  =  0.14679  ohm. 

An  approximate  rule  for  calculating  the  resistance  of  copper 
wire  per  thousand  feet,  in  ohms,  is  to  divide  10,000  by  the 


HAULAGE  COSTS 


273 


size  of  the  wire  in  circular  mils.  Thus,  for  a  4-0  wire  (211,600 
circ.mils),  the  resistance  is,  approximately, 

#=^^=0.04726  ohm  per  1000  ft. 

This  rule  should  only  be  used  in  rough  calculations.  When 
accuracy  is  desired,  the  resistance  for  the  wire  should  be  taken 
from  electrical  tables,  as  above  stated. 

The  resistance  of  steel  rails,  for  the  same  cross-section  and 
length,  varies  with  the  composition  of  the  steel.  The  presence 
of  sulphur  and  manganese,  particularly  the  latter,  greatly 
modifies  the  resistance  of  the  steel.  It  has  been  found  that 
this  resistance  will  vary  from  about  eight  to  thirteen  times 
that  of  copper  of  the  same  sectional  area  and  at  the  same 
temperature.  This  has  given  rise  to  what  is  termed  the  "ratio 
of  rail  to  copper"  or  the  " rail-to-copper  ratio. "  The  results 
of  numerous  experiments  have  made  it  possible  to  calculate  the 
equivalent  circular  mils  of  copper  corresponding  to  any  given 
weight  of  rail  in  pounds  per  yard,  for  any  rail-to-copper  ratio. 
To  do  this,  the  weight  of  rail,  in  pounds  per  yard,  is  multiplied 
by  the  constant  corresponding  to  this  ratio,  as  determined  by 
the  composition  of  the  steel.  The  value  of  this  constant,  for 
the  several  ratios,  is  as  follows: 


Rail-to-Copper 
Ratio 

Constant 

Rail-to-Copper 
Ratio 

Constant 

8 

15,550 

11 

11,360 

9 

13,820 

12 

10,360 

10 

12,500 

13 

9,590 

Applying  this  method  and  assuming  a  rail-to-copper  ratio 
of  10,  the  constant  for  this  ratio,  as  taken  from  the  above  table, 
is  12,500.  Then,  for  a  25-lb.  rail,  the  equivalent  circular  mils 
of  copper  is  25  X  12,500  =  312,500.  Electrical  tables  are  not 
generally  extended  to  include  as  large  a  wire  as  this  area 
indicates.  The  resistance  in  ohms  per  thousand  feet,  however, 
can  be  calculated,  approximately,  by  the  rule  previously  given. 
Thus,  10,000  -j-  312,500  =  0.032  ohm.  The  resistance  of  the  two 


274  COAL  MINING  COSTS 

rails,  in  the  first  2000  ft.  of  this  track,  is  the  same  as  the  resist- 
ance of  a  single  25-lb.  rail  1000  ft.  long,  or  0.032  ohm. 

Again,  for  a  30-lb.  rail  of  the  same  composition,  the  equiv- 
alent circular  mils  of  copper  is  30  X  12,500  =  375,000.  The 
corresponding  resistance  is,  therefore,  approximately,  10,000 
-f-  375,000  =  0.0267  ohm.  The  resistance  of  the  two  rails,  for 
1000  ft.  of  track,  is  one-half  of  this  amount,  or  0.01335  ohm. 
The  total  rail  resistance  in  this  track  is,  therefore,  0.032  -\- 
0.01335  =  0.04535  ohm ;  and  the  difference,  in  favor  of  the  iron 
rails,  is  0.14679  —  0.04535  =  0.10144  ohm. 

The  diagram,  Fig.  7,  taken  from  the  Ohio  Brass  Co.'s 
catalog,  shows  graphically  the  circular  mils  of  copper,  of  equal 
electrical  resistance  to  steel  rails  of  different  weights  and  "  rail- 
to-copper  ratios. "  The  curved  lines  show  the  resistance,  in 
microhms,  of  steel  rails  of  different  weights  and  ratios. 

The  diameter,  in  inches,  of  a  copper  wire  that  is  the  elec- 
trical equivalent  of  a  steel  rail  of  a  given  weight  (Ib.  per  yd.) 
may  be  calculated  by  multiplying  the  respective  constant  taken 
from  the  above  table,  by  the  weight  of  the  rail,  extracting  the 
square  root  of  the  product  and  dividing  that  result  by  1000 
Thus,  for  a  rail-to-copper  ratio  of  10,  the  constant  is  12,500. 
Then,  the  diameter  of  the  copper-wire  equivalent  is: 

V 25X12,500 


As  regards  the  second  problem,  a  30-hp.  motor  consumes 
30X746  =  22,380  watts  or  22.38  kw.  If  this  motor  is  taking,  as 
stated,  100  amp.,  the  voltage,  at  the  full  capacity  of  the  motor, 
is  22,380-^100  =  223.8  volts.  The  drop  in  voltage  for  this  line 
is,  therefore,  250-223.8  =  26.2  volts.  The  work  performed  by 
this  motor,  in  each  hour,  when  working  at  its  rated  capacity,  is 
22.38  kw.-hr. 

Tests  made  at  a  colliery  in  Pennsylvania  showed  that  the 
bonding  was  so  bad  that  the  return  current  was  leaving  the 
rails  and  finding  its  way  to  the  generating  plant  by  way  of 
the  ditches  and  water  pipes,  which,  of  course,  had  a  high  resist- 
ance. 

The  locomotives  with  a  rated  speed  of  6.3  miles  per  hour 
were  found  to  be  traveling  at  about  2.5  miles  and  making  an 
average  of  12  trips  per  day.  The  actual  load  on  the  generator 


HAULAGE  COSTS  275 

was  30  per  cent  over  the  rated  load  on  the  line.  The  main 
haulage  roads  were  bonded  with  compressed  terminal  bonds 
of  sufficient  capacity  to  equal  the  size  of  trolley  and  feeders. 

The  first  month  following  this  installation  ike  production 
was  the  largest  in  the  history  of  the  mine.  The  locomotives 
instead  of  making  12  trips  per  day  were  averaging  18  trips. 
Later  another  locomotive  was  added  and  the  whole  load  on 
the  generator  was  less  than  that  carried  before  the  bonding 
was  changed. 

At  another  colliery  the  bonding  resistance  of  the  main 
haulage  road  was  reduced  80  per  cent  by  the  use  of  compressed 
terminal  bonds.  During  the  last  six  months  of  the  channel- 
pin  installation  the  track  bonder  had  averaged  three  days  per 
week  on  this  road  replacing  broken  and  defective  bonds.  Dur- 
ing the  first  seven  months  of  the  new  installation  the  bonder 
spent  two  hours  on  the  road  replacing  some  bonds  broken  by 
a  wrecked  trip. 

Investigation  of  bad  haulage  conditions  at  another  mine 
disclosed  a  reading  on  the  voltmeter  of  180  volts  on  a  250-volt 
circuit  at  a  point  6000  ft.  from  the  mine  mouth.  A  10-ton 
motor  was  used  on  the  main  haulage  in  to  this  point,  beyond 
which  there  were  three  motors  engaged  on  secondary  haulage. 
When  the  main  haulage  motor  started  out  with  a  trip  the 
current  would  drop  to  80  volts  and  remain  at  this  for  8  to 
10  min.  during  which  time  the  gathering  motors  could  not 
move.  It  was  estimated  that  the  haulage  charges  at  this  mine 
were  increased  about  lOc.  per  car  or  4c.  per  ton  from  these 
causes  which  on  the  basis  of  a  working  schedule  of  20  days 
per  month  would  entail  a  loss  of  $800  per  month. 

TABLE  SHOWING  COST  OF  ENERGY-LOSS  IN  RETURN  CIRCUIT  PER  100 
AMPERES  LOAD,  PER  YEAR  OF  240  9-HouR  DAYS,  POWER  COSTING 
ONE  CENT  PER  KILOWATT-HOUR  (1913) 

One  Mile.  Two  42-lb.  rails  in  parallel,  assuming  continuous  joints. .  $15.78 
One  Mile.  Two  42-lb.  rails  in  parallel,  bonded  with  channel-pin 

bonds,  resistance  found  by  actual  test 37 . 80 

One  Mile.     Two  42-lb.  rails  in  parallel,  bonded  with  compressed  termi- 
nal bonds,  resistance  found  by  actual  test 17 . 99 

Fixed  cost  of  rail  resistance 15 . 78 

Increased  cost  with  compressed  bonds 2  21 

Increased  cost  with  channel-pin  bonds 22 . 02 

Saving  per  year  with  compressed  bonds  for  power  alone  per  100 

amperes 19.81 


276 


COAL  MINING  COSTS 


A  method  to  overcome  bonding  troubles  and  to  reduce  the 
bad  effects  to  a  minimum  is  to  put  in  ground  wires  along  the 
track  and  to  tap  these,  at  suitable  distances,  to  the  rails  as 
shown  in  Fig.  15;  this  insures  the  bonding  and  provides  an 
uninterrupted  metallic  circuit.  The  high  price  of  copper  prac- 
tically prohibits  its  use  for  this  purpose. 

Wornout  wire  hoisting  ropes,  in  lengths  of  500  ft.  and  over, 
which  make  few  joints  necessary,  having  a  scrap  value  of 
approximately  $8  per  ton  (as  of  1912)  and  a  specific  resistance 
of  8  to  1  as  compared  with  copper,  may  be  utilized  for  this 
purpose.  It  is  a  better  conductor  than  rails  and  a  rope  of, 


=: r:   EEI 


FIG.  15. — Method  of  improving  the  bonding' by  the  use  of  old  wire  cable. 

say,  ll/2  in.  diameter  has  about  the  same  current-carrying 
capacity  as  a  No.  000  B.  &  S.  gage,  or  about  a  %-in.  copper  wire. 
Smaller  ropes  may  also  be  used,  the  size  required  depending 
upon  the  conditions. 

The  wire  rope  may  also  be  used  where  ground  wires  are 
required,  as  between  substations  and  the  mines.  For  this  work, 
copper  cables  1  in.  in  diameter  are  quite  frequently  used.  An 
installation  requiring  1000  ft.  of  such  copper  cable  was  esti- 
mated at  approximately  $550  in  1912,  and  the  same  result 
would  be  obtained  with  wire  rope  at  a  cost  approximating 
$100.  For  such  outside  work  1%-in.  and  2-in.  ropes,  which 
cannot  be  'handled  very  well  inside,  may  be  utilized.  The 
installation  should  be  made  in  the  following  manner: 

From  the  substation  to  the  entrance  of  the  mine,  where  no 
rails  are  used,  1%-in.  or  2-in.  rope,  either  single  or  in  multiple, 


HAULAGE  COSTS  277 

according  to  the  load  on  the  circuit  and  the  distance,  should  be 
used.  Along  the  track  in  the  main  gangway  1%-in.  rope,  and 
for  extensions  1^4  and  1-in.  single  ropes  are  suitable.  The  rope 
is  stretched  out  by  attaching  it  to  a  mine  motor  or  locomotive 
and  in  this  manner  several  thousand  feet  may  be  laid  in  a 
few  hours.  As  the  ropes  are  seldom  shorter  than  500  ft.,  only 
a  small  number  of  joints,  which  should  be  made  with  sub- 
stantial 10-in.  clamps,  are  necessary.  Sheet-lead  lining  should 
be  used  between  the  clamps  and  rope,  thus  providing  a  reliable 
contact.  One  of  the  bolts  used  with  the  clamps  serves  as  a 
bond  terminal.  Where  the  rail  taps  make  connections  with 
the  ropes,  smaller  clamps  are  used  at  distances  of  about  200 
to  300  ft.  It  is  advisable,  in  order  to  increase  the  durability, 
to  put  the  rope  on  the  high  side  of  the  track  to  keep  it  out 
of  the  water  as  much  as  possible. 

The  efficiency  of  the  return  track,  either  single  or  double  ^Ki* 
bonded,  and  provided  with  such  a  ground  wire,  will  be  very 
little  affected  by  a  few  defective  bonds,  and  the  practical 
results  will  be  far  superior  than  with  the  regular  bonding 
only.  On  account  of  the  stable  conditions  obtained. with  the 
ground  wire,  in  many  cases  single  bonding  with  some  cross 
taps  between  the  rails  will  be  sufficient;  where  distances  are 
short  and  the  load  light,  satisfactory  results  will  be  attained 
by  the  ground  tapped  to  the  rails  at  suitable  distances  with- 
out the  regular  bonding.  The  inspections  of  the  track  in 
regard  to  bonding  will  also  be  reduced  to  a  minimum. 

At  one  operation  where  three  SV^-ton  electric  mine  loco- 
motives were  working  there  was  considerable  complaint  due 
to  the  inefficiency  of  these  motors  and  in  fact  a  request  was 
made  for  an  additional  motor.  An  examination  of  the  place 
showed  that  there  was  lost  time  for  power,  a  large  amount  of  sand 
was  being  used  on  the  track  and  trolley  wheels  did  not  last 
longer  than  a  day  or  two.  The  track  at  this  operation  was  laid 
with  40-lb.  rail  at  the  bottom  of  the  shaft  and  30-lb.  rail  in  the 
gangways.  The  bonding  was  about  equal  to  the  average  con- 
dition of  bonding  in  mines. 

As  promptly  as  possible,  7500  ft.  of  wire  rope  was  installed, 
2000  of  which  was  iy2-in.,  4000  114-in.  and  the  balance  1  in. 
in  diameter.  The  results  were  entirely  satisfactory.  The  motors 
could  then  do  good  work,  the  amount  of  sand  used  on  the 


278  COAL  MINING  COSTS 

track  has  been  reduced  65  per  cent,  which  also  reduces  to  a 
minimum  the  wear  on  the  motor  wheels.  There  is  no  further 
complaint  regarding  power  or  poor  trolley  wheels. 

Track  costs. — The  determining  factors  in  fixing  grades  in 
the  mines  are  drainage,  haulage  and  sometimes  loading.  The 
grade  sought  after  in  the  mines  of  this  country  is  4  to  6  in. 
per  100  ft.  There  is  a  theoretic  grade  at  which  the  inclination 
of  the  haulageway  will  compensate  the  added  resistance  due 
to  loading  the  cars  and  thus  equalize  the  load  on  the  motor 
for  both  empty  and  loaded  trips.  When  this  condition  has 
been  realized  the  motor  is  able  to  handle  the  same  number  of 
loads  and  empties  going  in  their  respective  directions. 

The  greater  the  difference  in  weight  between  the  loaded 
and  empty  car,  the  steeper  should  be  the  grade  in  favor  of  the 
load,  to  overcome  the  extra  resistance.  Consequently  when- 
ever a  larger  capacity  mine  car  is  adopted,  a  corresponding 
change  in  the  standard  grade  should  be  made  for  all  haulage- 
ways  not  influenced  by  other  considerations.  This,  however, 
is  seldom  done,  the  original  grade  being  maintained.  On  haul- 
ageways  where  the  0.35  per  cent  grade  is  used  with  cars  of  3 
tons  capacity,  the  number  of  empties  it  is  possible  for  the 
motor  to  haul  invariably  exceeds  the  number  of  loads. 

An  all-steel  car  (115  cu.  ft.  capacity  at  water  level  full) 
empty  weighs  5080  Ib.  and  loaded  with  coal  weighs  12,230  lb., 
while  loaded  with  rock  it  would  weigh  17,000  lb.  Assuming 
the  sum  of  the  resistance  due  to  friction  and  track  to  be  30  lb. 
per  ton  for  an  empty  car  and  25  lb.  per  ton  for  a  loaded  car, 
the  coefficient  of  friction  per  short  ton  for  loaded  and  empty 
cars  will  be  0.015  and  0.0125  respectively. 

With  an  8-ton  motor  having  a  rated  drawbar  pull  of  3000  lb. 
on  the  level,  letting  N  =  the  number  of  cars  and  g  =  the  theoretic 
grade  for  equalizing  the  drawbar  pull,  then  when  the  motor  is 
handling  coal-loaded  and  empty  cars: 

3,000+16,000g  3,000-16,000^ 

(12,230  X 0.0125)  - 12,2300  ~~  (5080 X 0.015) + 50800 
0  =  5  in.  per  100ft. 

With  the  same  motor  handling  empty  and  loaded  rock  cars: 

3000+16,0000  3000=16,0000 

17,000X0.0125-17,0000  =     5080X0.015+50800 
0  =  7  in.  (almost)  per  100  ft. 


HAULAGE  COSTS  279 

The  foregoing  is  a  somewhat  involved  quadratic  equation  not 
always  easy  of  solution.  A  method  which  is  perhaps  simpler  is 
as  follows: 

Let  6  =  desired  angle  of  grade  to  give  equal  drawbar  pull  in 
both  directions,  then  the  percentage  of  grade,  or  the  rise  in  unit 
length  of  track,  is  practically  equal  to  sin  6. 

If     W  —  weight  of  empty  car; 
W'  =  weight  of  loaded  car; 
C  =  coefficient  of  friction  for  empty  car; 
C'  ~  coefficient  of  friction  for  loaded  car; 

then  WC+  W  sin  6  =  W'C'  -  W  sin  8. 

Substituting  values  in  the  above  case 
(5080X0.015)  +5080  sin  0  =(12,230X0.0125)  -12,230  sin  0 
76.2+5,080  sin  0=152.875-12,230  sin  0  17,310  sin  0  =  76.675 

0.00443 


Multiplying  by  100  to  secure  the  rise  in  100  ft.,  we  have 
0.443  ft.,  or  about  51/4  in. 

With  straight  track,  then,  the  8-ton  motor  on  a  grade  of 
7  in.  per  100  ft.  could  pull  in  28  empty  cars  and  return  with 
the  same  number  loaded  with  rock;  but  as  the  resistance  of 
the  curves,  of  which  a  major  portion  of  every  gangway  con- 
sists, is  a  corollary  of  the  weight,  and  determined  grade  should 
be  augmented  somewhat  in  favor  of  the  loaded  cars  to  partly 
compensate  for  curve  resistance. 

The  value  of  the  track  resistance  per  ton  on  straight  track 
will  have  to  be  determined  experimentally  for  any  type  of  car. 
Equipping  the  cars  with  patent  bearings  will  have  salutary 
effects,  should  this  resistance  prove  too  high. 

On  a  recent  test  of  nine  cars,  three  each  of  the  three  various 
types  ordinarily  employed,  on  fairly  clean  straight  track,  on 
the  surface,  with  cars  weighing  12,230  Ib.  each;  loaded  with 
coal,  capacity  115  cu.  ft.  loaded  water  level;  wheel  base,  3  ft. 
6  in.  ;  gage  of  track  3  ft.  6  in.  ;  40-lb.  rail  ;  both  wheels  tight 
on  the  axle  ;  center  of  drawbar  pull  to  top  of  rail,  18%  in.,  the 
tractive  effort  required  was  as  follows: 

21.2  Ib.  per  ton  for  cars  with  160  deg.  iron  case,  babbitt 


COAL  MINING  COSTS 

lined;  10.5  Ib.  per  ton  for  cars  with  160  deg.  brass  lining; 
6.5  Ib.  per  ton  for  cars  with  Hyatt  bearings. 

The  cars  were  of  all-steel  construction  identical  in  every 
respect,  except  the  bearings.  The  brass-bearing  cars  were  new 
and  in  first-class  condition,  while  the  Hyatt  and  babbitt-bear- 
ing cars  had  been  in  use  almost  2  yr.  The  results,  if  secured 
on  tracks  underground,  in  their  usual  dirty  condition  and  with 
inferior  ballast,  would  no  doubt  have  been  higher. 

A  contingency  not  always  considered  in  planning  a  haulage 
is  that,  since  the  abandonment  of  mule  haulage,  the  loaders 
are  compelled  to  move  the  cars  at  the  loading  chutes  by  their 
own  exertions.  Four  cars  per  trip  are  not  infrequently  loaded 
from  one  chute.  On  the  prevalent  0.35  per  cent  grade  to  move 
the  cars  the  distance  necessary  to  accomplish  this  would  require 
an  additional  man.  To  dispense  with  this,  the  foreman  resorts 
to  increasing  the  grade  immediately  beneath  the  chutes  with 
an  equalizing  diminution  between  them.  This  expedient  per- 
mits loading  the  cars,  but  makes  the  roadbed  a  series  of  undula- 
tions with  pools  of  water  in  the  "dead  spots."  It  raises  the 
haulage  cost,  destroys  the  rolling  stock  and  sets  the  cars 
bumping  and  jerking  over  tlie  entire  working  section.  It  is 
needless  to  mention  that  a  heavier  grade  would  abate,  if  not 
wholly  remove,  this  condition. 

To  be  in  congruity  with  the  preceding,  a  grade  of  7  to  8  in. 
per  100  ft.  for  the  haulage  and  up  to  8^  in.  for  the  primary 
gangways  would  seem  to  have  more  to  recommend  it  than 
the  lesser  grade.  To  be  sure,  the  first  installation  ,of  the 
heavier  grade  would  reduce  the  available  lift  as  the  gangway 
advanced,  but  this  objection  would  remedy  itself  in  all  sub- 
sequent levels. 

Again,  a  ditch  averaging  18  in.  wide  edge  to  edge  and  6  in. 
deep  in  the  center,  on  a  0.35  per  cent  grade,  would  give  38.4  ft. 
velocity  and  almost  108  gal.  capacity  per  minute,  against 
54.4  ft.  and  153  gal.  for  an  8%-in.  grade. 

If  an  analysis  of  any  haulage  problem  indicates  that  a 
saving  is  possible  by  cutting  down  grades,  it  becomes  at  once 
necessary  to  determine  the  approximate  amount  that  can  be 
profitably  spent  on  the  proposed  improvement. 

The  advantages  accruing  from  the  improvement  will  be: 
(1)  Reduction  in  general  labor  costs;  (2)  reduction  of  power 


HAULAGE  COSTS  281 

cost  per  ton  hauled;  (3)  reduction  in  general  expense,  includ- 
ing the  saving  made  possible  by  the  postponement,  either  tem- 
porarily or  permanently,  of  expenditures  for  additional  equip- 
ment such  as  motors,  generating  units,  engines  and  boilers. 

If  a  haulage  motor  pulls  a  certain  number  of  extra  cars 
per  trip  as  a  result  of  a  grade  reduction,  the  total  number  of 
tons  produced  at  the  mine  will  be  increased  without  a  pro- 
portionate increase  in  the  expense.  This  saving  is  expressed 
by  a  formula  in  which  let : 

t  =  tons  produced  per  shift  before  increasing  the  tonnage; 
jP=tons  produced  per  shift  after  increasing  the  tonnage  by 

pulling  extra  cars  per  trip; 
c  —  labor  cost  to  produce  t  tons,  in  dollars; 
C  =  labor  cost  to  produce  T  tons,  in  dollars; 
S  =  saving  per  ton  in  dollars  resulting  from  the  increased 

tonnage. 
We  then  have  the  formula: 

„    c    C 

s=t~r 

For  example,  a  mine  with  one  main-line  haulage  motor 
making  16  trips  in  8  hr.  from  two  partings  produces  950  tons 
per  day  at  a  cost  of  $145.38  for  inside  labor  and  $52.69  for 
outside,  making  a  total  of  $198.07. 

Investigation  of  the  haulage  profile  shows  that  by  reducing 
the  grade  on  one  of  the  runs  the  motor  will  be  able  to  handle 
1000  tons  per  day.  To  handle  this  increased  tonnage  an  extra 
driver  inside  will  be  required  and  a  track  layer,  one  hour  per 
shift,  bringing  the  total  labor  cost  including  full  allowance  for 
feed,  car  and  depreciation  of  the  mule  up  to  $202.45.  Sub- 
tituting  these  values  in  the  above  formula,  we  have : 

198.07     202.45 
"950" "TOOO" 

"When  the  summit  in  a  grade  is  lowered,  the  power  required 
to  overcome  grade  resistance  is  less  and  by  increasing  the 
number  of  cars  per  trip  less  power  is  required  per  ton  of  coal 
hauled  since  the  proportionate  ton-mileage  of  the  motor  itself 
is  reduced.  The  weight  of  the  motor  inbound  may  be  about 


282  COAL  MINING  COSTS 

one-third  of  the  total  weight  of  the  trip,  and  outbound  about 
one-fifth,  so  that  it  is  evident  that  considerable  power  is  con- 
sumed in  moving  it  alone. 

A  heavy  pull  exerted  on  a  stiff  grade  will  tend  to  increase 
maintenance  and  repair  charges  for  both  rolling  stock  and 
track.  On  the  other  hand  a  reduction  in  the  grade  may  result 
in  an  increase  in  the  car  repair  bill,  due  to  the  greater  num- 
ber of  cars  handled,  but  the  charges  per  car-mile  and  the 
maintenance  per  ton  hauled  will  remain  the  same  and  will  not 
effect  the  unit  cost  per  ton  of  coal  produced. 

The  elimination  of  bad  grades  also  reduces  the  hazard  to 
operations  and  though  the  danger  to  life  is  not  directly  cal- 
culable, it  is  of  vital  importance  and  must  not  be  overlooked 
in  considering  the  possibility  of  any  proposed  improvement; 
in  other  words  it  is  simply  applied  safety  and  could  properly 
be  included  under  the  charge  for  insurance. 

If  the  advantages  resulting  from  a  certain  grade  elimination 
can  be  reduced  to  cents  per  ton  handled  over  the  section  of 
track  improved  and  this  is  multiplied  by  the  number  of  tons 
so  handled,  the  result  will  be  the  amount  that  may  be  expended 
on  the  proposed  work.  Or  expressed  in  a  formula,  let: 

2/=the  estimated  number  of  tons  available; 

C  =  cost  of  grade  reduction,  including  interest  on  money 

invested; 

S  =  summation  of  all  savings,  in  cents  per  ton; 
V  =  value  of  safe  operation. 

Then  (yS  +  V)  should  be  equal  to,  or  greater  than,  the 
value  of  C.  If  the  value  of  yS  (total  saving  in  dollars)  is  less 
than  the  cost  of  the  improvement,  and  the  value  of  safe  opera- 
tion does  not,  in  the  opinion  of  the  management  overcome  the 
difference,  then  the  project  should  be  abandoned.  As  a  matter 
of  fact  the  judgment  of  the  financier  must  be  relied  upon 
throughout  the  whole  study  of  the  question.  False  assumptions 
may  be  made  in  some  cases,  leading  to  erroneous  conclusions, 
but  by  following  carefully  the  steps  indicated,  it  is  possible 
to  mafce  a  fairly  accurate  estimate  of  the  results  to  be  obtained 
in  any  contemplated  work  of  this  description. 

Haulage  grading  estimates. — When  all  the  data  for  a  pros- 
pective change  of  grades  has  been  assembled,  the  estimated 


HAULAGE  COSTS 


283 


cost  of  the  various  plans,  routings  and  schemes  should  be 
made  for  purposes  of  comparison.  An  intelligent  estimate  of 
cost  must  consider  detail  and  be  based  on  accurate  knowledge 
of  the  proposed  requirements,  together  with  the  application  of 
the  unit  costs  of  similar  work  formerly  completed.  A  con- 


FORM  No.l 

ABC  COAL  COMPAls 

General  Estimate                                                                                     Propose 

IY 

d  Extension  of  Motor  Road 

Pn,H~a 

Ft.  anil  Construction  pf                                      ft. 

MINE  N<-» 

ITEM 

Tool                      Rail* 
Laid 
Taken  up 

35*; 

°ii*"i** 

"wliki** 

prti 

*cS" 

air 

* 

Tou                      Rail. 
UN 

Taken  op 

Kegs  Small  Spikes 

Site 

Kegs  Motor  Spikes 
Site 

Site 

Sm.ll  Tlea 

Pn   FUB  Plate* 
No.  Rail* 
No.  Rails 

Ken  Track  Bolu 
No.  Rail* 
No.  Rail* 

Bono* 
.         Bonding  Cap* 
Bonding  Sleeves 

Trot*  anil  Switege* 
No.  Rail* 
No.  Rails 

2|o  Trolley  Wlrs 
4|0,  Trolley  Wire 

Hancor*  and  Clamps 
Wire  Splicing,  Sleeve* 
Trolley  Fret* 
Automatic  Cut-out  Swltcasr 
Trolley  wire  Guard* 
Intulated  Telephone  Wire 
Porcelain  Insulator*  and  Pitt 
Material  to  os  DeliTsrsd 
Cleaning  Oob.  Etc. 

_ 

Griding..      Yd*.  Top 
Griding.    .        .'.Yd*.  Bottom. 

Setting  Props 
Setunc  Timber  Sets  and  Cross  Ban 

Mlae.  Eixnie 

Total  Eitlmited  COM 

Credits  of  Material  not  applied  19  estimate. 

Small   Tit, 

Total  Value  Credits  to  be  Deducted 
Grand  Total  FstlmAtiJ  Cost 

The  extension  of  this  Motor  Road  will 

AftMftsffe 

JffMMrffe 

FIG.  16. — Form  for  assembling  estimate  of  cost  for  grade  revision  in  a 

motor  road. 

venient  form  for  assembling  an  estimate  of  cost  of  a  motor  road 
extension  or  revision  is  shown  in  the  form  Fig.  16.  The  detail 
required  is  not  exacting,  yet  a  fairly  complete  record  is  indi- 
cated. When  it  is  finally  decided  to  carry  through  the  work  a 
final  estimate  is  made  and  a  copy  bearing  the  official  signature 


284 


COAL  MINING  COSTS 


of  approval  sent  forward  to  all  departments  concerned.    The 
record  is  thus  made  complete. 

Economy  in  the  execution  of  the  work  will  depend  on  the 
evolution  of  a  systematized  method  and  a  strict  adherence 
to  two  fundamental  principles  of  good  management :  First  the 

FORM  No.  2 


DAILY  TIME  AND  PROGRESS  CHART 
MOTOR  ROAD  CONSTRUCTION 


Name 


Cefvn  Amoll 


E2 


Ma  nws  Hgyr 


MATERIAL 
RECEIVED 


PROGRESS  CHART 

Note:  Place  a  cross  n  -the  squareslo  indicate  the  portion  completed  between  each  station 


Cleanin 


Takin    u    Track 


Drillin    and  Shoot  n 


Loadinqjfock 


Drivers  or  Motorman 


Unloadinqjtock 


3onding  &,  Hanging  Wir 


Timberin    and  Misc. 


XXX 


XIX 


Correct   ^n  tAmitfi,   Foreman- 
Checked    «Wi  £vam>    Supt. 

FIG.  17. — Progress  chart  for  use  in  analyzing  grade  revision  costs. 

number  of  men  employed  in  any  one  section  in  any  period  shall 
be  adjusted  to  the  amount  and  classification  of  the  work  to  be 
done  and,  secondly,  only  experienced  men  should  be  employed. 
A  careful  study  of  the  progress  made  from  day  to  day 
will  plainly  show  any  weak  points  in  the  organization  and 
will  often  indicate  a  probable  remedy.  Progress  charts,  coupled 


HAULAGE  COSTS  285 

with  the  report  of  the  daily  time,  are  useful  in  judging  the 
comparative  efficiency  of  the  organization  as  a  whole  or  in 
part  and  also  gives  a  record  of  the  unit  costs  for  making 
future  estimates  and  comparisons.  The  accompanying  form, 
Fig.  17,  shows  a  good  method  of  accomplishing  this.  The 
progress  for  any  particular  section  can  be  indicated  in  the 
daily  report  by  filling  in  the  squares  of  the  progress  report 
at  the  bottom  of  the  form  with  crosses  to  express  the  estimated 
amount  of  work  completed  in  that  section.  Thus  a  daily  chart 
can  be  sent  to  the  administrative  department  as  a  record  of 
both  the  work  done  and  the  efficiency  of  the  organization. 

Haulage  track  curves. — Viewed  solely  from  the  haulage 
standpoint,  the  determining  factors  of  the  curve  radius  can 
be  covered  by  two  heads: 

1.  The   cost  of  resistance   due  to  curvature  on  the  total 
estimated  number  of  cars  that  can  be  hauled. 

2.  The  probable  number  of  cars  to  be  hauled  in  each  trip 
and  the  speed  of  haulage. 

The  amount  of  resistance  due  to  curvature  varies  with  each 
type  of  car,  and  to  a  lesser  degree  with  each  car  of  a  certain  type. 
This  resistance  expressed  in  terms  of  grade,  with  curves  of  from 
30  to  100  ft.  radius,  will  run  0.015  ft.  to  0.025  ft.  per  100  ft.  of 
track  for  each  degree  of  curvature.  That  is,  with  a  50-ft.  radius, 
or  115-deg.  curve,  moderately  clean  track,  fair  running  cars  with 
both  wheels  keyed  on  the  axle,  approximately  a  1.8  per  cent  down 
grade  would  be  necessary  to  secure  the  same  drawbar  pull  as  on 
a  tangent.  The  value  of  a  curve  expressed  in  degrees  can  be 
obtained  by  dividing  5730  by  the  radius  in  feet.  This  formula 
will  have  to  be  employed  especially  in  small  radius  curves;  the 
actual  arc  is  used  to  find  the  degree,  rather  than  the  central  angle 
subtending  the  100-ft.  cord,  the  practice  on  standard-gage  roads. 

By  using  the  actual  arc,  a  50-ft.  radius  =   _     =115  deg.  curve; 

ou 

50 

by  using  a  100-ft.  cord,  a  50-ft.  radius  =   .    1,  =  180  deg.  curve, 

sin  -^oL 

showing  a  disparity  of  65  deg. 

Assuming  a  curve  with  a  central  angle  of  90  deg.  a  ruling 
grade  of  0.5  per  cent  and  allowing  the  same  rate  of  resistance 
per  degree  on  a  25-ft.  and  50-ft.  radius  curve,  the  motor  in 
traveling  over  the  two  would  have  to  work  equivalent  to 


286  COAL  MINING  COSTS 

mounting  a  4.5  per  cent  grade  for  39  ft.  and  a  2.5  per  cent 
grade  for  78  ft.  respectively.  From  the  beginning  of  the  50-ft. 
radius  to  the  point  of  tangent  there  would  be  a  total  of  1.96  ft. 
^\  vertical,  while  to  travel  between  the  same  points  by  way  of 
the  25-ft.  radius  curve,  including  the  25  ft.  of  tangent  on  each 
end  of  the  curve,  there  would  be  a  total  of  2.02  ft.  vertical, 
or  essentially  the  same  vertical  rise  in  either  case. 

While  actually  with  the  smaller  radius  curve  there  would 
be  a  lower  rate  of  resistance  per  degree,  this  would  be  more 
than  balanced  by  the  increased  resistance  due  to  the  slower 
speed  compelled  by  the  sharper  curve.  If  the  resistance  due 
to  grade  and  curvature  between  the  similarly  located  points  is 
accepted  as  equal,  then  there  remain  in  favor  of  the  50-ft. 
radius  the  greater  speed  at  which  the  trip  can  travel,  the 
reduced  danger  from  cars  jumping  the  track,  the  better  adher- 
ence of  trolley  to  wire,  and  11  ft.  shorter  haul,  and  under 
«-%  A  some  conditions  11  ft.  less  of  track.  With  a  gangway  produc- 
ing six  trips  per  day  of  twelve  5-ton  cars  each,  this  11  ft. 
twice  per  trip  would  consume  enough  power  to  draw  1  ton 
7920  ft.  each  day,  or  375  mi.  per  year.  A  self-recording 
dynamometer  will  reveal  the  frictional  resistance  of  any  type 
of  car,  and  the  monetary  value  per  unit  of  haulage  can  be 
readily  ascertained. 

In  estimating  the  number  of  cars  per  trip,  the  future  out- 
put, as  well  as  the  length  of  haul,  must  be  considered.  A 
haulage  over  which  100  cars  travel  per  day  may  be  increased 
threefold  by  a  tunnel  to  the  veins.  This  will  mean  the  instal- 
lation of  a  larger  motor  if  the  haul  is  long,  or  possibly  the 
use  of  two  motors.  The  maximum  speed  of  haulage  under- 
ground being  fixed  by  law  (6  mi.  per  hour),  nothing  can  be 
expected  from  faster  transportation.  A  15-ton  locomotive  will 
not  traverse  curves  possible  to  an  8-ton  machine,  and  this 
fact  will  demand  an  extra  motor,  with  its  attendant  expense. 

For  obvious  reasons  no  compensation  is  allowed  for  curva- 
ture underground;  and  if  a  motor  is  required  to  work  at  its 
capacity,  the  additional  resistance  to  be  overcome,  because  of 
curvature,  will  be  the  factor  limiting  the  length  of  the  trip. 
With  the  large  curve  a  locomotive  may  pull  through  on  its 
potential  velocity,  but  on  a  curve  of  25  or  30  ft.  radius,  the 
velocity  will  have  to  be  reduced  before  reaching  the  curve. 


HAULAGE  COSTS 


287 


Rails. — Much  extra  expense  can  be  incurred  by  not  having 
the  weight  of  rail  used  in  the  track  properly  proportioned  to 
the  weight  of  the  motor  operating  on  it.  Where  the  rail  is 
too  heavy,  there  is  an  unnecessary  expenditure  in  first  cost 
and  where  too  light,  as  is  more  frequently  the  case,  costs  of 
track  maintenance,  together  with  extra  wear  and  tear  on 
motors,  due  to  poor  track  conditions,  will  mount  up  rapidly, 
though  perhaps  not  be  so  evident.  Working  under  average 
conditions,  the  Baldwin-Westinghouse  Co.  recommend  the  fol- 
lowing minimum  weight  rails  for  general  mine  service  with 
motor  haulage : 


Weight  of  Motor 
in  Tons 

Weight  of  Rail  in 
Pounds  per  Yard 

4  to    6 

16 

6  to    8 

20 

8  to  10 

25 

10  to  13 

30 

13  to  15 

40 

15  to  20 

50 

Mines  having  average  size  mine  cars  customarily  use  40  to 
60  Ib.  rail  on  the  main  haulage,  20  to  40  Ib.  on  secondary 
haulage  and  16  to  25  Ib.  in  the  rooms. 

An  approximate  rule  sometimes  used  for  this  purpose  is 
to  have  a  rail  that  will  weigh  at  least  4  Ib.  per  yard  for  each 
ton  of  weight  in  the  locomotive.  For  example,  a  4-ton  motor 
should  run  on  a  16-lb.  rail;  a  5-ton  on  a  20-lb.  rail,  etc.  This 
rule  gives  somewhat  excessive  rail  weights  when  applied  to 
the  heavier  types  of  motors  and  the  above  table  is  preferable 
if  available. 

In  purchasing  rail  for  mine  use  the  buyer  will  require: 
(1)  Stiffness,  (2)  strength  and  (3)  durability  rather  than  tons 
of  steel.  If  the  strength  of  various  sections  is  compared,  it 
will  be  found  that  these  requisites  can  be  purchased  at  a  lower 
unit  rate  in  the  larger  sections.  In  " stiffness"  we  have  that 
property  which  allows  the  rail  to  span  the  ties  and  support 
the  load  without  bending,  affording  thereby  a  smooth  running 
surface  for  the  cars;  in  "strength"  we  have  that  quality  which 


288  COAL  MINING  COSTS 

bears  the  load  without  yielding  or  breaking,  while  in  "dura- 
bility ' '  we  have  the  ability  to  resist  wear  over  extended  periods 
of  time. 

The  stiffness  varies  as  the  square  of  the  weight,  and  the 
strength  as  the  3/2  power,  while  the  price  per  ton  is  nearly 
constant.  If  the  unit  weight  is  assumed  as  being  30  Ib.  per  yd., 
then  the  stiffness  will  increase  as  follows: 

THIRTY  POUNDS  PER  YARD — STIFFNESS  =  1 

16f  per  cent  increase  in  weight  35  Ib.  per  yard  stiffness  =  1 . 36  or  a   36  per  cent 

increase. 
33 1  per  cent  increase  in  weight  40  Ib.  per  yard  stiffness  =  1 . 78  or  a  78  per  cent 

increase. 
50  per  cent  increase  in  weight  45  Ib.  per  yard  stiffness  =  2 . 25  or  a  125  per  cent 

increase. 
66f  per  cent  increase  in  weight  50  Ib.  per  yard  stiffness  =  2 . 79  or  a  179  per  cent 

increase. 
100  per  cent  increase  in  weight  60  Ib.  per  yard  stiffness  =  4 . 00  or  a  300  per  cent 

increase. 

The  ultimate  strength  will  increase  as  follows: 

THIRTY  POUNDS  PER  YARD — ULTIMATE  STRENGTH  =  1 

16f  per  cent  increase  in  weight  35  Ib.  per  yard  ultimate  strength  =  1 . 26  or  a 

26  per  cent  increase. 
33  £  per  cent  increase  in  weight  40  Ib.  per  yard  ultimate  strength  =  1 . 54  or  a 

54  per  cent  increase. 
50    per  cent  increase  in  weight  45  Ib.  per  yard  ultimate  strength  =  1 . 84  or  a 

84  per  cent  increase. 
66f  per  cent  increase  in  weight  50  Ib.  per  yard  ultimate  strength  =  2. 15  or  a 

115  per  cent  increase. 
100    per  cent  increase  in  weight  60  Ib.  per  yard  ultimate  strength  =  2 . 83  or  a 

183  per  cent  increase. 

The  advantages  of  the  heavy  section  over  the  light,  as 
regards  stiffness  and  strength,  would  show  a  higher  comparison 
as  the  rail  wears  or  wastes  away  from  any  cause  whatsoever. 

In  determining  the  durability  of  rail,  it  is  obvious  that  a 
great  amount  of  wear  cannot  be  expected  if  the  weight  selected 
conforms  closely  to  the  immediate  duty  it  has  to  withstand. 

We  can  assume  for  practical  purposes  that  half  the  total 
weight  is  in  the  head,  and  that  about  half  of  this  weight,  or 
one-quarter  the  weight  of  the  rail,  can  be  worn  away  before 
the  rail  is  discarded,  if  a  sufficient  margin  of  metal  has  been 


HAULAGE  COSTS 


289 


allowed;  otherwise,  the  rail  will  fail  before  it  has  attained 
much  more  than  a  high  polish. 

In  mining  work,  particularly  underground,  with  the  track- 
men in  absolute  charge,  trackwork,  derailments,  rail  breakage, 
etc.,  are  taken  as  part  of  the  day's  routine  and  pass  unnoticed, 
except  that  part  which  appears  indirectly  in  the  high  main- 
tenance charges. 

If  we  assume  that  a  wear  of  1/5  the  weight  of  the  head 
was  allowed  as  a  safety  factor  in  the  lighter  rail,  then  the 
durability  of  light  and  heavy  sections  will  compare  as  follows : 


AVAILABLE  FOR  WEAB 

Spare 

Times 

Increase 

Metal  in 

Increase 

in 

Left  in 
Head 

Next 

of  Wear 

Weight 

Weight 
in 
Pounds 
per  Yard 

Weight 
in 
Head 
Only 

Maximum 
Half 
Head 

Minimum 
One-fifth 
Head 

After 
Minimum 
Wear 

Heaviest 
Rail  before 
Head 
Becomes 
as  Light 

by 
Adding 
5Lbs. 
to 
Section 

by 
Adding 
5Lbs. 
to 
Section 

30 

15.0 

7.5 

3.0 

12 

5.5 

1.830 

1/6 

35 

17.5 

8.75 

3.5 

14 

6.0 

1.710 

1/7 

40 

20.0 

10.00 

4.0 

16 

6.5 

.625 

1/8 

45 

22.5 

11.25 

4.5 

18 

7.0 

.550 

1/9 

50 

25.0 

12.50 

5.0 

20 

7.5 

.500 

1/10 

55 

27.5 

13.25 

5.5 

22 

8.0 

.454 

1/11 

60 

30.0 

15.00 

6.0 

24 

8.5 

.420 

1/12 

Or,  using  30-lb.  rail  as  a  unit,  the  metal  available  for  wear 
would  compare  as  follows: 


AVAILABLE  FOR  WEAR  BEFORE 

Weight 
in  Pound 

Weight 
in  Head 

HEAD  WOULD  BECOME  AS  LIGHT 

Increase 
in  Weight 

per  Yard 

Only 

Maximum 

Minimum 

per  Yard 
Per  Cent 

Per  Cent 

Per  Cent 

30 

15 

7.  5  or  100 

3.0  or  100 

35 

17| 

10.0  or  133| 

5.  5  or  183! 

16f 

40 

20 

12.  5  or  166| 

8.0  or  266| 

33| 

45 

22! 

15.0  or  200 

10.  5  or  350 

50 

50 

25 

17.  5  or  233| 

13.0  or  433! 

66f 

55 

27* 

20.0  or  266| 

15.5or516f 

83! 

60 

30 

22.  5  or  300 

18.0  or  600 

100 

290 


COAL  MINING  COSTS 


Briefly,  if  we  were  about  to  build  a  permanent  (so-called) 
narrow-gage  road  for  mine  traffic,  for  which  30-lb.  steel  would 
ordinarily  be  used,  we  would  gain,  by  using  a  60-lb.  section, 
the  economy  in  maintenance,  a  more  easily  operated  road  with 
its  attendant  benefits,  fewer  ties,  fewer  derailments  and  a 
larger  scrap  value  when  the  rail  was  reclaimed.  Furthermore, 
we  would  have  a  stiffness  four  times,  an  ultimate  strength  2.83 
times  and  a  durability  three  to  six  times  as  great,  for  a  rail 
expenditure  but  double  that  for  30-lb.  steel. 

Some  concerns,  by  purchasing  "second"  rail  from  the  rail- 
road companies,  obtain  the  heavier  rail  for  the  same  price  per 
lineal  foot  as  for  new  sections  one-half  to  two-thirds  their 
weight.  This  quality  of  rail  for  most  mining  purposes  will 
serve  as  well  as  new  sections. 

In  localities  where  acid  water  abounds  the  corroding  of  the 
steel  is  frequently  the  limiting  factor  in  the  life  of  the  rail. 
It  would  be  futile  to  lay  heavy  section  rail  in  locations  where 
the  water  would  soon  destroy  it.  As  the  web  and  edges  of 
the  flange  are  the  portions  destroyed  first,  an  inspection  of 
the  standard  dimensions  will  evidence  that  by  increasing  the 
weight  we  do  not  secure  a  proportionate  increase  in  the  acid- 
resisting  properties  of  the  rail.  Rail  weighing  25  Ib.  per  yd. 
has  been  taken  as  the  basis  of  unity. 


Weight 
of  Rail 

Increase 
in  Weight, 
Per  Cent 

Thickness 
of 
Web 

Increase  in 
Thickness, 
Per  Cent 

Thickness 
Ends  of 
Flange 

Increase  in 
Thickness, 
Per  Cent 

25 

it 

... 

H 

30 

20 

ft 

11 

ft 

35 

40 

H 

21 

« 

9 

40 

60 

it 

32 

H 

27 

45 

80 

H 

42 

H 

35 

50 

100 

If 

47 

H 

36 

60 

140 

ft 

63 

H 

64 

In  the  standard  tee  rail,  adopted  by  the  American  Society 
of  Civil  Engineers,  42  per  cent  of  the  metal  is  in  the  head, 
21  per  cent  in  the  web  and  37  per  cent  in  the  flange.  The  top 
corners  are  curved  to  a  5/iG-in-  radius,  and  the  car  wheels  are 
designed  to  give  on  this  as  little  friction  as  possible;  as  the 


HAULAGE  COSTS 


291 


rail  more  nearly  wears  to  the  shape  of  the  flange  the  friction 
is  augmented.  The  height  of  the  rail  is  identical  with  the 
width  of  the  flange,  so  if  this  dimension  is  measured  the  weight 
can  be  determined. 

The  table  shows  the  weight  of  rail  per  yard  corresponding 
to  the  height  of  flange  width. 

Track  frogs. — Standardization  of  switches  and  frogs  at 
mines  to  a  limited  number  of  sizes  to  meet  requirements  will 
substantially  lower  the  cost  of  making  these.  The  accompany- 
ing illustration,  Fig.  18,  shows  a  standard  frog  and  switch 
used  by  the  O'Gara  Coal  Co.  in  1916. 


iof7Je. 
•OB 
End  Elevation 


Side     Elevation 
Switch 


FIG.  18. — Standard  frog  and  switch  used  by  the  O'Gara  Coal  Co. 

The  designs  were  made  after  considering  both  simplicity 
and  economy  of  construction.  Any  ordinary  blacksmith  or 
ironworker  will  make  these  parts  without  difficulty.  The  cost 
of  making  a  No.  5  frog  and  two  6-ft.  switch  points  at  a  well- 
equipped  mine  shop  was  about  1916:  Material,  $6.37;  labor 
of  blacksmith  and  machinists,  $6.58 ;  total,  $12.95. 

A  cast-steel  frog  supplied  by  manufacturers  at  a  cost  of 
about  $6.50  is  inherently  more  rigid  than  the  riveted  frog,  but 
it  is  difficult  to  fasten  it  securely  to  the  ties.  In  order  to  stiffen 
the  riveted  structure,  cast-iron  fillers  may  be  added  which  also 
support  the  flange  of  the  wheels  in  passing  over  the  throat  of 
the  frog,  thus  relieving  the  jar  to  the  rolling  stock. 


292 


COAL  MINING  COSTS 


DIMENSIONS  OF  STANDARD  FROGS  AND  SWITCHES  FOR  NARROW-GAGE 
INDUSTRIAL  AND  MINE  TRACKS 

Standard  Frog  for  Motor  Turnout  (Right  or  Left) 


Frog 
No. 

Frog 
Angle, 
X 

Rail, 
per 
Yard 

Length 
of  Frog, 
A 

Wing 
Rail, 
B 

Heel 
Distance, 
C 

Length  of 
Throat, 
D 

Straight 
Rail, 
E 

Deg.  Min. 

Pound 

Ft.    In. 

In. 

Ft.    In. 

In. 

Ft.    In. 

3 

18      55 

30 

4      0 

16 

2      8 

3i| 

2      3 

4 

14       15 

30 

4      8 

20 

3      0 

5| 

2      9 

4 

14       15 

40 

4      8 

20 

3      0 

6| 

3      0 

5 

11       25 

30 

4     10 

20 

3      2 

7& 

3       0 

5 

11       25 

40 

5      0 

20 

3      4 

81 

3       0 

Standard  Switch  for  Motor  Turnout  (Right  or  Left) 


Weight 

Length 

Distance 

Length 

Length 

of 

of 

between 

of 

Rail 

Gage 

of 

Rod 

Rail, 

Point, 

Bridle 

Rail 

Punching, 

of 

Bridle 

Punch- 

Pound 

F 

Rods 

Planed, 

Track, 

Rod, 

ing, 

per 

Ft.     In. 

L 

H 

J     K 

G 

M 

N 

Yard 

Ft.     In. 

Ft.     In. 

Ft.     In. 

In. 

Ft.     In. 

In. 

20 

4      0 

(1  rod) 

1       4 

4      2 

36 

5         SI 

25 

25 

4      0 

(1  rod) 

1       6 

4       2 

40 

6        Oi 

29 

30 

4      0 

(1  rod) 

1       9 

4      2 

42 

6        21 

31 

30 

6      0 

2      3 

2       6 

4      2 

44 

6        4| 

33 

30 

7      6 

3      0 

3       0 

4      2 

48 

6        8| 

37 

40 

6      0 

2      3 

2      9 

5      2i 

40 

7      6 

3      0 

3       6 

5       2| 

The  throw  of  switch  point  is  3^  in.  all  for  cases. 

In  designing  various  parts  of  the  turnouts  it  was  kept  in 
mind  that  all  such  turnouts  may  be  of  only  temporary  useful- 
ness in  one  particular  location  and  that  the  constituent  parts 
may  be  used  many  times  before  being  cast  aside  as  useless. 
The  standard  frog  is  somewhat  shorter  than  one  designed  to 
the  specifications '  of  the  American  Railway  Engineering  As- 
sociation, but  the  saving  in  weight  and  bulk,  with  the  conse- 
quent saving  in  making  the  several  installations,  will  more 
than  offset  any  loss  due  to  instability. 


HAULAGE  COSTS 


293 


The  cost  of  laying  and  ballasting  a  No.  5  turnout  complete, 
as  shown  in  Fig.  19,  was  about  1916  as  follows : 

One  30-lb.  No.  5  frog  and  two  6-ft.  points $12.95 

40  ties,  5X6  in.,  at  20  cts 8.00 

Spikes,  bolts,  tie-plates,  etc .75 

1  low  switch  stand  and  rods 2 . 25 

2  headblocks,  5  X6  in.,  8  ft.  long 1 .00 


Total  material $24.95 

Laying,  16  hr.  at  35?  cts 5.68 

Ballasting  and  surfacing,  8  hr.  at  35|  cts 2 . 84 

Total  labor $8.52 

Total  cost  of  material  and  labor $33 . 47 

K- 


FIG.  19. — Standard  turnout  used  by  the  O'Gara  Coal  Co. 

The  dimensions  of  the  standard  frogs,  switches  and  turn- 
outs are  given  in  the  accompanying  tables.  The  formulas  used 
for  the  turnouts  are  as  follows : 

G-B  sin  X-F  sin  Y 


Chord  length  U-- 
Radius  R- 


sin%  (X+Y) 
G-B  sin  X-F  sin  Y 


-K7; 


cos  Y— cos  X 
Lead  S=(R+%G)(sin  X-sin  Y) 

+BcosX+F+0   ; 
in  which 

X=frog  angle; 

F  =  angle  of  point  rail; 

B= length  of  wing  rail; 

F  =  length  of  switch  rail; 

0  =  distance  from  actual  to  theoretical  frog  point: 

G  =  gage  of  track; 

R  =  radius  of  turnout. 


294 


COAL  MINING  COSTS 


The  dimension  of  0  was  taken  as  2  in.  and  the  heel  distance 
of  switch  points  as  4^4  in. 

The  spacing  of  the  ties  depends  on  the  size  of  the  tie  and 
the  style  of  the  turnouts.  If  the  regular  set  of  switch  ties  is 
used  as  in  standard-gage  trackwork,  5  X  6-in.  ties  spaced  18  in. 
center  to  center  will  give  good  results  for  track  laid  with 
rails  of  up  to  40  Ib.  in  weight.  If  the  turnout  is  laid  with  ties 
of  even  length  staggered  in  as  shown  in  Fig.  19,  a  spacing  of 
16  to  18  in.  for  each  branch  has  proved  satisfactory.  This 
style  of  construction  is  specially  well  adapted  to  underground 
turnouts,  where  headroom  is  limited  and  flat  ties  3  X  5  in.  or 
3  X  6  in.  are  used.  All  switch  ties  should  be  of  hardwood  and 
treated  if  possible,  as  decay  will  set  in  before  mechanical  wear 
destroys  their  usefulness. 

DIMENSIONS  FOR  TURNOUTS  IN  NARROW-GAGE  TRACKS 
Gage  of  Track,  36  In. 


Frog 
No. 

Length 
of 
Switch 
Points, 

Radius  of 
Turnout, 

Length  of 
Lead, 

Length  of 
Straight 
Rail, 

Chord 
Curved 
Rail, 

Mid- 
ordinate 
Curved 
Rail, 

F 

R 

S 

T 

U 

V 

Ft. 

Ft.      In. 

Ft.      In. 

Ft.      In. 

Ft.      In. 

In. 

3 

4 

42     7& 

16      3H 

10      9H 

10     10& 

4& 

4 

4 

81     1| 

19      3 

13      5 

13       8 

3| 

4 

6 

75    8£ 

22       6 

14      8 

14     lOf 

4& 

5 

4 

141     71 

22      2f 

16      41 

16       7A 

2| 

5 

6 

126    7 

26      0| 

18      2| 

18       41 

m 

5 

71 

122    91 

28      4£ 

19      0£ 

19       2| 

4& 

6 

n 

184    9f 

30      9f 

21       5| 

22      9 

4A 

Gage  of  Track,  42  In. 


3 

4 

52  Iff 

18   91 

12  11| 

13   31 

4H 

4 

4 

99  2| 

22   3 

16   5 

16   8| 

4| 

4 

6 

92  6| 

25   9 

17  11 

18   2| 

6J 

5 

4 

172  0| 

25   9| 

19  lit 

20   2 

31 

5 

6 

153  8f 

29  11| 

22   \\ 

22   31 

4H 

5 

71 

149  If 

32   5| 

23   \\ 

23   4 

5^ 

6 

7£ 

223  51 

36   1\ 

27   3| 

27   6 

5^ 

Track. — The  following  is  an  interesting  example  of  estimat- 
ing the  cost  of  laying  5000  ft.  of  track  where  the  grade  is  1  per 


HAULAGE  COSTS  295 

cent  in  favor  of  the  loads ;  a  12-ton  motor  is  used  and  3*/2  ton 
cars  assuming  that  the  rails  cost  $26  per  ton,  ties  lOc.  each, 
spikes  $3.75  per  keg  of  200  lb.,  labor  for  trackmen  $2.50  per 
day,  and  helpers  $1.75  per  day,  these  figures  being  as  of  1911. 

The  rails  for  a  12-ton  motor  haulage  should  not  be  lighter 
than  40  lb.  per  yard,  which  would  require  (2  X  5000  X  40) 
-f-  (3  X  2240)  =  59.5  tons  at  a  cost  of  26  X  59.5  =  $1547.  For 
40  lb.  rails,  use  3V2  X  Vie  in.  spikes,  12  kegs,  at  $3.75  per  keg 
=  $45 ;  and  4  X  6  in.  cross-ties,  spaced  2  ft.  center  to  center, 
2500  at  lOc.  each  =  $250.  There  will  be  required  also,  using 
24-ft.  rails,  832  angle  or  fish-plates,  6240  lb.,  at  iy2c.  per  pound 
—  $93.60,  and  8  kegs  bolts,  nuts,  and  washers  at  $5  per  keg  = 
$40;  making  the  total  track  material  $1975.60.  The  laying 
and  surfacing  of  5000  ft.  of  track  in  mine  entries,  under  ordi- 
nary conditions,  including  the  handling  of  the  material  in  the 
shaft  and  its  distribution  in  the  entry  will  require,  approxi- 
mately 120  days'  labor  for  helpers  at  $1.75  per  day  =  $210; 
50  days,  trackmen,  at  $2.50  per  day  =  $125 ;  and  6  days, 
drivers,  at  $2  per  day  =  $12 ;  total  for  labor  $347.  The  total 
cost  of  the  track  laid  is  therefore  $2322.60,  making  no  allow- 
ance for  special  grading  which  might  be  required  at  some  points 
in  the  entry. 

A  number  of  interesting  figures  on  the  comparative  cost  of 
track  laid  with  steel  and  wood  mine  ties  under  varying  con- 
ditions were  given  in  a  paper  presented  before  the  West 
Virginia  Mining  Institute,  in  1913  from  which  the  following 
have  been  excerpted: 

The  Peyton  Block  Coal  Co.  used  four  steel  mine  ties  per 
rail  length  which  at  a  cost  of  32c.  each  made  a  total  of  $1.28 
per  each  pair  of  rails.  Under  the  same  conditions,  11  wood 
ties  would  be  required,  which  at  a  cost  of  5c.  and  allowing  lOc. 
for  spikes  brought  the  total  cost  per  pair  of  rails  to  65c.  Off- 
setting this  difference,  however,  it  was  found  that  the  miners 
would  lay  track  with  steel  ties  in  their  working  places  them- 
selves while  they  always  insisted  on  the  regular  mine  track 
layer  performing  the  work  when  wood  ties  were  used  because 
of  the  special  tools  and  labor  required.  It  was  believed  that 
this  saving  in  the  time  of  the  track  layer  compensated  for 
the  difference  in  the  cost  of  the  material  involved,  so  that  the 
extra  life  of  the  steel  tie  could  be  regarded  as  clear  profit. 

The  Allegheny  River  Mining  Co.  advances  the  opinion  that 


296 


COAL  MINING  COSTS 


the  life  of  the  steel  mine  tie  is  about  six  times  that  of  the 
wood  tie  and  since  only  one-half  as  many  are  required  per  foot 
of  track,  the  ratio  of  comparative  utility  was  1  to  12.  On  the 
basis  of  32c.  for  the  steel  ties  and  8c.  for  wood,  which  was  the 
delivered  cost  to  this  company,  and  disregarding  cost  of  the 
spikes  required  for  wooden  ties,  the  cost  ratio  is  4  to  1. 

Track  costs  at  the  Dartmore  Mine  of  the  Davis  Coal  & 
Coke  Co.,  when  laid  with  steel  ties  were  found  to  be  $1.28  per 
30-ft.  length  for  material.  When  using  wooden  ties,  10  of  these 
were  required  per  30-ft.  length  of  track  which,  at  a  cost  of  lOc. 
each  and  allowing  12c.  for  spikes,  brought  the  cost  of  material 
for  the  track  with  wooden  ties  up  to  $1.28  per  30  ft.  The 
difference  in  the  cost  of  material  for  wooden  and  steel  ties  at 
this  mine  amounted  to  %c.  per  lineal  foot  of  track,  disregard- 
ing economies  in  laying,  salvage,  etc.  as  indicated  above. 

At  mines  Nos.  14  and  20  of  the  same  company,  it  was  found 
that  the  steel  tie  saved  sufficient  head  room  to  eliminate  the 
necessity  of  brushing  the  top  and  effecting  a  computed  saving 
of  63c.  per  yard  of  track. 

The  Hutchinson  Coal  Co.  compiled  an  interesting  study  of 
the  comparative  cost  of  steel  and  wooden  ties  at  its  Kirkwood 
Mine,  near  Bridgeport,  Ohio,  during  the  years  1907  to  1909 
inclusive  when  it  was  changing  from  the  wood  to  the  steel 
ties.  In  the  1907  period  all  wood  ties  were  being  used;  in 
the  1908  period  75  per  cent  steel  ties  were  in  use  and  in 
1909  all  steel  ties  were  in  use.  The  comparative  figures  are  as 
follows : 


19 

07 

19 

08 

19 

09 

Output 

Cost  in 
Cents 
per  Ton 

Output 

Cost  in 
Cents 
per  Ton 

Output 

Cost  in 
Cents 
per  Ton 

September.  . 

16,083 

4  80 

12,620 

3  98 

21,556 

2.84 

October  

26,216 

4.02 

12,620 

3.36 

22,540 

2.81 

November  

22,617 

3  70 

16,281 

3.21 

25,191 

2.82 

Average    cost   per 
ton 

4  10 

3.48 

2.82 

HAULAGE  COSTS  297 

The  saving  per  ton  with  all-steel  track  thus  appears  to  be 
1.28c.,  which  in  the  three  months  period  of  1909  amounted  to 
$886.85. 

This  same  company  also  compiled  an  interesting  compara- 
tive estimate  of  the  cost  of  track  in  a  room,  computed  on 
the  basis  of  a  width  of  24  ft.,  length  of  200  ft.  and  a  thickness 
of  coal  of  5  ft.  4  in.  which  worked  out  as  follows : 


WITH  WOOD  TIES 

Ties,  1\  ft.  apart,  80  ties  at  12  cts $9 . 60 

320  spikes  equals  40  Ib 90 

Laying  and  removing  track,  labor 7 . 50 

Depreciation  of  ties  and  spikes 3 . 50 


Total $21.50 

Salvage 7.00 

Net  cost..  .  $14.50 


WITH  STEEL  TIES 

35  ties  at  33  cts $11 . 55 

Labor,  removing  (track  laid  by  the  miners) 1 . 50 

Depreciation 1 . 00 


Total $14.05 

Salvage 10. 55 

Net  cost..  $3.50 


Estimating  the  output  of  coal  from  this  room  at  1000  tons 
on  which  a  saving  of  $11  will  be  effected,  this  amounts  to 
l.lc.  per  ton. 

After  an  exhaustive  series  of  tests  the  Carnegie  Steel  Co. 
prepared  the  accompanying  comparative  cost  of  track  laid  in 
rooms  with  steel  and  wooden  ties.  The  table  is  based  on  rooms 
280  ft.  long  with  steel  ties  spaced  at  4  ft.  center  to  center  and 
wood  ties  2  ft.  center  to  center.  This  estimate  contemplates 
using  each  wood  tie  in  two  consecutive  rooms  and  after  the 
second  year  renewing  annually  15  per  cent  of  steel  ties,  or  10 
new  ties  per  year,  per  room. 


298 


COAL  MINING  COSTS 


Number  of  Ties  in  One  Room 

Steel,  70 

Wood,  140 

Cost  of  ties  f.o.b.  mine  in  carload  lots  
560  spikes 

@  0.29      $20.30 

@  0.06       $8.40 
@  0  00|       2  80 

@  0  02     $1  40 

@  0  04         5  60 

@  0  01           70 

@  0  02         2  80 

Maintenance  cost  for  first  room  steel  ties 

$2  10         2  .  10 

Total  cost  at  end  of  life  of  first  room 

$22  .  40 

$19  60 

@  0  02     $1  40 

@  0  04         5  60 

560  spikes                             

@  0.00$       2  80 

Taking  up  when  room  worked  out  
Maintenance  cost  for  second  room  steel  ties 

@  0.01          .70 
$2  .  10         2  .  10 

Total  cost  at  end  of  life  of  second  room 

$24  .  50 

$28  00 

Saving   per  room  in  favor  of 
steel  ties             $3  .  50 

@  0.29     $2.90 

@  0  05           50 

$2.40 

@  0.06       $8  40 

560  spikes                           

@  0.00^       2  80 

@  0  02     $1  40 

@  0  04         5  60 

@  0  01           70 

@  0  02         2  80 

$4  50       $4  50 

Total  cost  at  end  of  life  of  third  room  

$29  .  00 

$47  .  60 

Saving   per   room  in  favor  of 
steel  ties                                        $18  60 

$2  40 

560  spikes                   

@  0.00£       2.80 

@  0  02     $1  40 

@  0.04         5.60 

Taking  up  when  room  worked  out  

@  0.01          .70 

Maintenance  cost  for  fourth  room  steel  ties  . 

$4  .  50       $4  .  50 

Total  cost  at  end  of  life  of  fourth  room  .  . 

$33  .  50 

$56.00 

Saving   per  room  in  favor  of 
steel  ties    $22  .  50 

The  matter  of  track  friction  is  important,  and  most  mining 
men  realize  that  there  are  material  advantages  in  a  good  track. 
Few,  however,  really  comprehend  the  reduction  in  power 
requirements  that  can  be  effected  on  much-used  roads  by  mak- 
ing a  strictly  firstclass  track  in  every  respect.  When  we  con- 
sider that  it  is  possible  to  have  a  track  friction  as  low  as 
12  Ib.  of  drawbar  pull  per  ton,  while  as  many  roads  show  as 
much  as  40  Ib.  per  ton,  it  is  apparent  that  this  is  an  extremely 


HAULAGE  COSTS 


299 


important  item.    The  use  of  heavy  rails  is  an  essential  feature, 
but  it  is  even  more  necessary  that  the  track  be  kept  clean. 

Mine  cars. — In  1911  a  wooden  car  of  49  cu.  ft.  capacity, 
without  brakes,  cost  at  a  certain  mine  in  the  neighborhood  of 
$45  each.  The  same  company  was  offered  steel  cars  of  the 
same  outside  dimensions,  but  having  greater  capacity,  as  fol- 
lows: 


Capacity 

Cost,  Delivered 

Increase,  Per  Cent 

7  cu.  ft. 

$65.00 

45 

1  cu.  ft. 

62.50 

38 

1  cu.  ft. 

75.00 

66 

5  cu.  ft. 

68.50 

52 

The  cost  of  an  all-steel  car  thus  ran  from  38  to  66  per  cent 
more  than  a  wooden  one,  in  some  cases  more,  possibly,  espe- 
cially in  case  of  the  addition  of  improvements  in  the  way  of 
draft  gear,  running  gear  or  wheels.  The  manufacturers  them- 
selves state  the  increase  of  cost  to  be  from  50  to  100  per  cent. 

Steel  cars  have  a  somewhat  greater  capacity  than  the 
wooden  car,  varying  somewhat  with  the  design  of  the  car, 
but  ranging  generally  between  10  and  20  per  cent.  The  accom- 
panying table  gives  the  comparative  capacities  and  weights  of 
some  steel  and  wooden  cars  of  the  same  dimensions,  taken 
from  actual  practice: 


CAPACITY,  CUBIC  FEET 


WEIGHT  IN  POUNDS 


Wood 

Steel 

Increase, 
Per  Cent 

Wood 

Steel 

Per  Cent 
Gain  in 
Capacity 

Saving 
in  Weight, 
Per  Cent 

49 

58 

18 

2010 

1800 

18.0 

11.5 

94 

105 

111 

53 

60 

13 

1920 

1985 

13.2 

3.0* 

14 

19 

35 

1085 

855 

35.0 

26.0 

*  Increase 


300  COAL  MINING  COSTS 

Weights  of  both  wooden  and  steel  cars  vary  widely  and 
it  is  difficult  to  make  any  accurate  comparison.  Steel  cars 
are  sometimes  heavier  than  the  wooden  ones  of  the  same 
capacity,  but  generally  they  appear  to  run  from  10  to  20  per 
cent  less  in  weight  for  the  same  capacity. 

The  chief  advantages  of  the  steel  car  are : 

That  although  the  first  cost  of  the  steel  car  is  greater,  the 
increased  life  and  decreased  cost  of  maintenance,  together 
with  increased  capacity,  thereby  necessitating  a  fewer  number 
of  cars  to  handle  a  given  output,  more  than  make  up  for  the 
difference. 

That  the  advantage  of  the  steel  car  having  a  greater 
capacity  with  the  same  outside  dimensions,  or  the  same  capacity 
with  smaller  dimensions,  is  of  great  value,  especially  in  low 
veins. 

The  increased  capacity  of  the  steel  car  should  materially 
reduce  the  cost  of  haulage,  and  incidentally  tend  to  increase 
the  output  of  the  miner. 

The  saving  it  is  possible  to  effect  in  the  tare  weight  of  the 
car  itself  would  also  be  a  factor  in  the  reduction  of  costs  by 
reducing  the  proportion  of  dead  to  live  load. 

The  steel  cars  will  not  warp,  shrink  or  split,  which  are 
advantages  that  are  apparent  to  all,  besides  preventing  the 
leaking  of  dust  coal  on  haulways — not  only  a  nuisance,  but 
a  constant  source  of  danger  and  expense. 

On  a  special  type  of  steel  car,  repair  charges  have  been 
estimated  at  Ic.  per  50  ton-miles.  The  total  ton-miles  on 
which  this  was  computed  amounted  to  120,000  on  which  the 
repair  charges  were  $22.50. 

The  manufacturers  claim  that  the  cost  of  maintenance  of 
steel  cars  is  only  about  20  to  25  per  cent  of  that  for  wooden 
cars.  From  the  experience  of  the  railroad  companies  with 
their  steel-car  equipment,  it  would  seem  this  cannot  be  regarded 
as  an  underestimation.  Four  different  manufacturers  estimate 
the  life  of  the  steel  car  to  be  two  to  four  times  the  life  of  the 
wooden  cars. 

The  relation  of  the  tare,  or  weight  of  the  car,  to  its  capacity 
is  an  important  consideration  in  the  economics  of  haulage, 
the  tare  of  course  representing  dead  load  haulage.  The  accom- 
panying table,  compiled  from  bulletins  of  the  Illinois  Coal 


HAULAGE  COSTS 


301 


Mining  Institute,  and  data  received  from  car  manufacturers 
give  average  ratio  of  capacity  to  tare  for  the  normal  mine 
car. 


LIGHT  CARS 

MEDIUM  CARS 

Tare 

Capacity  of 
Cars 
in  Pounds 

Ratio  of 
Capacity 
to  Tare 

Tare 

Capacity  of 
Cars 
in  Pounds 

Ratio  of 
Capacity 
to  Tare 

800 
900 
1000 
1100 
1200 
1300 
1400 
1500 
1600 
1700 
1750 
1900 
2000 

Averag 

2000 
3000 
4200 
2600 
2600 
4200 
3000 
3500 
3000 
5000 
5600 
4000 
5000 

e  ratio  

72-28 
77-23 
80-20 
70-30 
68-32 
76-24 
68-32 
70-30 
65-35 
75-25 
76-24 
68-32 
72-28 

2200 
2400 
2525 
2665 
2850 

Averag 

6000 
6000 
4000 
4800 
4800 

e  ratio..  

73-27 
71-29 
61-39 
64-36 
63-37 

66-34 

HEAVY  CARS 

3240 
3330 
3500 
3700 
3780 

Averag 

6500 
6000 
6500 
8000 
6750 

e  ratio  

67-33 
65-35 
65-35 
68-32 
64-36 

72-28 

66-34 

At  mines  having  cars  already  equipped  with  plain  bear- 
ings which  it  is  desired  to  change  to  roller  bearings  the  cost 
under  average  conditions  will  be  about  $50  per  car,  this  figure 
being  as  of  1917.  Computing  on  the  basis  of  500  cars  this 
would  involve  an  expenditure  of  $25,000  less  the  salvage  value 
of  the  old  axles,  journal  boxes,  and  wheels  which  would  be 
about  as  follows  (figures  as  of  1917)  : 


1000  axles,  100  Ib.  each, 
2000  boxes,  25  Ib.  each, 
2000  wheels,  135  Ib.  each, 


$20  per  ton $1000 

$15  per  ton 375 

$15  per  ton 2025 


$3400 


The  actual  first  cost  of  the  roller-bearing  installation  would 
then  be  $25,000  —  $3400  =  $21,600.  The  interest  on  this  invest- 
ment is  $21,600  X  6  per  cent  =  $1296  per  year. 


302  COAL  MINING  COSTS 

Let  it  now  be  assumed  that  we  can  wipe  out  the  initial  invest- 
ment by  creating  a  sinking  fund.  To  provide  a  sinking  fund  for 
$1000,  for  example,  over  a  period  of  10  years,  which  is  the  con- 
servative life  of  the  bearings  if  properly  used,  we  must  lay  aside 
$75.87  each  year.  This  is  based  on  an  interest  rate  of  6  per  cent. 
For  $21,600,  we  must  lay  aside  $1638.79  each  year.  The  interest 
on  $21,600  and  the  sinking  fund  amounts  yearly  to  $1638.79  + 
$1296  =  $2934.79. 

Let  us  calculate  what  this  saving  amounts  to  in  the  case  of 
500  cars. 

Power  costs,  say  Ic.  per  ton  of  coal  hauled.  Drawbar  pull 
per  ton  (plain  bearings)  equals  32  Ib.  -|-  20  Ib.  for  every  1  per 
cent  of  grade.  Drawbar  pull  per  ton  with  roller  bearings 
equals  13  Ib.  +  20  Ib.  for  every  1  per  cent  of  grade.  These 
figures  are  from  dynamometer  tests  made  by  P.  B.  Lieber- 
mann,  chief  engineer  of  the  Hyatt  Roller  Bearing  Co. 

The  saving  in  drawbar  pull  equals  52  Ib.  —  33  Ib.  =  19  Ib. 
per  ton  hauled  up  a  1  per  cent  grade.  The  saving  in  power 
=  if  X  Ic.  =  0.4  mill  per  ton  hauled. 

Each  car,  let  us  assume,  hauls  6  tons  per  day.  The  saving 
in  power  per  year  with  500  roller  bearing  cars  =  0.4  mill  X 
6  tons  X  300  days  X  500  cars  =  $3600. 

Plain-bearing  cars  require  oil  once  a  day.  It  takes  two 
men  at  least  to  attend  to  the  oiling.  The  cost  of  oil  and  waste 
for  500  plain-bearing  cars  per  year  equals  $525.  Cost  of  two 
men  per  year  equals  $1200.  Total,  $1725.  The  cost  of  oil  and 
labor  for  500  roller-bearing  cars  is  $290.  The  saving  is  thus 
$1435. 

These  figures  are  the  results  of  carefully  made  tests.  With 
plain-bearing  cars  it  is  necessary  to  use  two  men  at  the  tipple, 
in  order  to  push  the  cars.  Since  roller-bearing  cars  push  with 
one-half  the  effort  required  for  plain-bearing  cars,  it  takes  but 
one  man  to  handle  the  roller-bearing  cars  at  the  tipple.  For 
the  same  reason,  the  services  of  one  eager  can  be  dispensed 
with  at  the  foot  of  the  shaft. 

The  saving  on  these  two  men  equals  $1875  per  year,  figur- 
ing that  the  man  at  the  bottom  costs  $2.75  per  day  and  the 
one  at  the  top  costs  $3.50  per  day. 

The  total  yearly  saving  on  500  roller-bearing  cars  is  thus 


HAULAGE  COSTS  303 

estimated  at:    Power,  $3600;  lubrication,  $1435;  labor,  $1875. 
Less  a  yearly  cost  of  $2944.79,  or  $3965.21. 

The  Hyatt  Roller  Bearing  Co.  conducted  experiments  with 
a  dynamometer  car  to  determine  accurately  the  actual  train 
resistance  of  cars  equipped  with  their  type  of  roller  bearing 
and  those  equipped  with  plain  bearings. 

One  of  these  tests  conducted  at  Greensburg,  Pa.,  on  a  track 
with  an  average  grade  showed  an  average  drawbar-pull  of 
12.8  Ib.  at  a  speed  of  5.98  miles  per  hour  for  the  roller  bearing 
and  24.3  Ib.  at  5.59  miles  per  hour  for  the  plain  bearing.  The 
advantage  in  favor  of  the  roller  bearings  works  out  at  47.25 
per  cent. 

A  second  test  at  Carbondale,  Pa.,  on  an  average  grade  of 
0.45  per  cent  showed  an  average  drawbar-pull  of  13  Ib.  at 
7.8  miles  per  hour  for  the  roller  bearing  as  compared  with 
32  Ib.  at  8  miles  per  hour  for  the  plain  bearing.  The  saving 
in  drawbar-pull  in  this  case  amounts  to  59.3  per  cent. 

The  diameter  of  the  car  wheel  has  an  important  influence 
on  the  power  required  to  move  the  car.  The  smaller  the  wheel 
the  more  difficult  it  is  to  move.  Cars  move  with  less  power 
on  the  narrower  track  gages  as  well,  other  conditions  being 
equal.  Wheel  bases  on  mine  cars  rarely  exceed  42  in.  and  the 
shorter  this  is  the  sharper  the  curve  the  car  will  negotiate. 

On  a  good  clean,  level  track  a  man  can  push  a  car  thai 
weighs  iy2  t°ns  an(i  carrying  a  load  of  2y2  tons,  making  a 
total  load  of  4  tons,  though  the  car  will  be  difficult  to  start  and  , 
stop.  It  is  probably  inadvisable  to  have  cars  that  have  to  be 
handled  by  men  weigh  when  loaded  over  3  tons  and  then  the 
track  should  be  in  good  condition. 

Many  progressive  operators  are  now  providing  a  stretcher 
car  for  emergency  purposes,  a  move  that  will  be  commended 
by  all  who  have  ever  had  occasion  to  assist  in  bringing  a  badly 
injured  man  from  the  mine.  Fig.  20,  shows  an  excellent  type 
of  car  for  this  purpose,  the  cost  of  which  was  $80,  including 
material  and  labor,  in  1916.  The  design  is  quite  simple,  being 
an  ordinary  mine  truck  with  the  stretcher  box  supported  on 
carriage  springs.  A  box  is  slung  from  the  truck  portion  of 
the  car.  In  this  will  be  kept  bandages,  salves  and  stimulants 
and  a  lungmotor.  The  car  will  be  kept  in  a  dry  place  specially 


304 


COAL  MINING  COSTS 


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HAULAGE  COSTS  305 

provided  for  the  purpose.     This  place  will  be  at  some  point 
near  the  face  of  the  workings. 

The  length  over  all  is  8  ft.,  the  width  2  ft.  8  in.,  and  the 
height  of  the  box  12  in.    The  bill  of  material  is  as  follows : 


Lumber  : 

2  pieces  4  X  6  in.  by  8  ft.  long,  yellow  pine. 

2  pieces  2  X  12  in.  by  7  ft.  8  in.  long,  yellow  pine. 

1  piece  2  X  10  in.  by  7  ft.  8  in.  long,  yellow  pine. 
4  pieces  2  X  6  in.  by  2  ft.  8  in.  long,  yellow  pine. 
4  pieces  2  X  8  in.  by  2  ft.  8  in.  long,  oak. 

2  pieces  2  X  4  in.  by  8  ft.  long,  yellow  pine. 

4  pieces  1  X  12  in.  by  7  ft.  8  in.  long,  yellow  pine. 
2  pieces  1  X  12  in.  by  2  ft.  8  in.  long,  yellow  pine. 
1  piece  1  X  10  in.  by  7  ft.  8  in.  long,  yellow  pine. 


Iron  and  bolts: 

60  %  X  1%-in.  carriage  bolts. 

40  l/2  X  4%-in.  carriage  bolts. 

35  1/2  X  9-in.  carriage  bolts. 

20  y2  X  3-in.  carriage  bolts. 

10  %  X  3-in.  carriage  bolts. 

1  iron  %  X  4  in-  X  8  ft. 

2  irons  %  X  4  in.  X  20  ft. 
10  irons  %  X  1%  in.  X  14  ft. 

1  %-in.  chain,  21  in.  long. 

2  irons  i/2  X  2-in.  X  30  ft. 
10  irons  %  X  1%-in.  X  14  ft. 

2  irons  %  X  2-in.  X  30  ft. 

5  %-in.  cut  washers. 
10  %-in.  cut  washers. 

1  1-in.  pin.,  8  in.  long,  with  12  in.  of  %-in.  chain  attached, 
to  couple  to  4  common  two-leaf  buggy  springs  made  of 
%  X  1%-in.  spring  steel;  length  over  all  3  ft.  with  8  in. 
between  springs. 

1  set  14-in.  car  wheels. 

2  2-in.  standard  axles,  3  ft.  1  in.  gage. 


306  COAL  MINING  COSTS 

Rope  haulage. — The  Chicosa  Fuel  Co.  in  Colorado  installed 
a  rope  haulage  system  about  1910,  operated  by  a  number  of 
small  electric  hoists.  Single-drum  hoists  are  used  on  the  cross- 
entries  and  double-drum  tail-rope  engines  on  the  levels. 

The  pitch  on  the  cross-entries  averages  81/3  per  cent,  the 
highest  is  13%  per  cent.  No  trouble  has  been  experienced  in 
hauling  6  cars  per  trip,  and  at  times  8  cars  per  trip.  The  car 
and  coal  combined  weigh  about  4900  Ib.  The  speed  of  the 
single-drum  hoists  working  on  the  cross-entries  is  300  ft.  per 
min.  on  empty  drum.  This  is  what  the  manufacturer  calls  the 
starting  speed.  The  300  ft.  per  min.  was  considered  the  best 
speed  for  switching  purposes,  after  several  speeds  had  been 
tried.  On  the  double-drum  hoists,  from  12  to  24  cars  are 
hauled  with  the  rope  speed  of  400  ft.  per  min. 

One  engineer,  one  rope  rider  and  one  hoist  can  move  as 
much  coal  as  8  or  10  drivers  and  mules,  besides  eliminating 
risks  of  killing  both  mules  and  drivers.  It  is  a  well-known 
fact  that  the  depreciation  on  machinery  is  much  less  than  that 
on  mules.  What  it  takes  to  feed  three  mules  will  pay  one 
engineer,  and  what  it  costs  to  keep  up  two  sets  of  mine  harness 
will  keep  in  repair  an  electric  hoist. 

There  are  at  present  five  50-hp.,  single-drum,  electric  hoists 
and  two  25-hp.,  double  drum,  tail-rope  hoists  in  operation. 
The  five  single-drum  hoists  work  on  cross-entries,  and  the  two 
double-drum  hoists  haul  the  coal  from  the  crosses  to  the  main- 
slope  partings.  The  entire  output  of  the  mines  which  amounts 
to  1100  tons  of  coal  per  10  hr.  is  moved  with  this  machinery. 

None  of  the  hoists  are  handling  at  present  more  than  50 
per  cent  of  their  capacity,  as  the  equipment  inside  the  mine 
could  easily  handle  2200  tons  of  coal  per  10  hr.  With  the 
present  output  the  cost  of  hauling  coal  inside  the  mine  is  35 
per  cent  cheaper  than  if  hauled  by  mules,  and  15  per  cent 
cheaper  than  if  hauled  by  mine  locomotives.  On  the  present 
basis  of  the  output  the  cost  per  ton  inside  the  mine  for  hauling 
from  the  rooms  to  the  main  slope  partings  is  3c.  per  ton;  by 
increasing  the  output  to  2200  tons  the  cost  would  l%c.  per 
ton.  For  each  1000  ft.  the  level  entries  are  driven  in,  the  cost 
will  only  increase  3  per  cent.  These  figures  are  based  on 
actual  tests  which  are  made  daily,  and  all  expenses,  labor, 
oil,  power,  interest  on  investment,  depreciation,  maintenance 


HAULAGE  COSTS  307 

of  hoists,  power  lines,  ropes,  and  bell  wire  system,  are  taken 
into  account. 

The  electric-haulage  system  which  is  now  being  used  is  far 
ahead  of  the  electric  locomotive.  Even  where  a  room-gather- 
ing locomotive  is  used,  the  expense  of  keeping  up  trolley  wires 
for  running  locomotives  is  70  per  cent  higher  than  keeping 
up  power  wires  in  back  entries.  The  danger  to  both  men 
and  mine  from  trolley  wires  is  100  per  cent  greater  than  from 
power  wires  for  electric  hoists.  The  troubles,  expenses,  and 
power  losses  in  rail  bonding  for  electric  locomotives  are  entirely 
done  away  with  in  the  electric  hoist,  and  the  cost  of  keeping 
up  the  electric  locomotive  will  be  more  than  keeping  up  the 
electric  hoist  and  rope. 

With  electric  locomotives,  heavy  steel  rails  are  necessary 
in  the  tracks,  while  ordinary  20-lb.  steel  rails  answer  the 
purpose  where'  the  electric  hoist  is  used.  This  item  makes  a 
difference  of  50  per  cent  in  track  cost  in  favor  of  the  electric 
hoist. 

The  electric  locomotives  cannot  be  used  in  gassy  mines, 
while  the  electric  hoist  can  be  used  with  perfect  safety. 

Electric  locomotives  cannot  be  used  to  an  advantage  on 
grades  exceeding  5  per  cent,  while  with  the  hoist  the  grade 
makes  no  difference. 

Gravity  planes. — Some  interesting  cost  figures  on  gravity 
plane  haulage  were  contained  in  a  paper  presented  before  the 
"West  Virginia  Mining  Institute  in  1914.  The  figures  applied 
to  an  installation  at  the  mines  of  the  LaFollette  Coal,  Iron 
&  Ry.  Co.  in  Tennessee. 

The  plane  operates  through  a  vertical  head  of  613  ft.  and 
a  horizontal  distance  of  3640  ft.  and  the  average  inclination 
is  16.8  per  cent.  It  is  operated  by  seven  men,  two  men  and 
the  drumman  at  the  head  house  to  attach  the  trips  and  two 
at  the  tipple  to  attach  the  empties  with  two  others  on  tho 
plane  to  test  the  grips,  oil  rollers,  etc. 

The  rope  for  the  plane  lasted  two  years  and  cost  $2000. 
In  1913  the  plane  handled  140,497  tons  in  268  days  of  9  hr.  or 
an  average  of  524  tons  per  shift.  The  maximum  tonnage 
handled  in  one  day  was  885  tons  in  &y2  hr.  which  is  at  the 
rate  of  104  tons  per  hour. 

It  will  be  seen  from  the  above  that  the  depreciation  on  the 


308  COAL  MINING  COSTS 

rope  amounts  to  0.7c.  per  ton  and  the  labor  of  the  seven  men, 
which  includes  oiling  the  cars,  amounts  to  $11.83  per  shift  or 
2i/4c.  per  ton.  Maintenance  costs  for  new  ties,  sheaves,  grips, 
etc.,  was  found  to  average  $25  per  month  or  0.2c  per  ton  of 
coal  handled;  this  includes  cleaning  up  wrecks,  replacing 
derailed  cars,  etc.,  there  being  an  average  of  one  wreck  per 
month  which  took  about  two  hours  to  clean  up.  The  total  cost 
of  operating  the  plane  in  cents  per  ton,  is:  Depreciation  on 
rope,  O.TOc. ;  labor,  2.25c. ;  maintenance,  0.20c. ;  total,  3.15c. 

Endless  rope  haulage. — The  expression  face  haulage,  indi- 
cates a  means  of  disposing  of  the  output  from  the  working 
face  by  a  haulage  system  sufficiently  flexible  to  be  capable  of 
rapid  extensions.  A  system  conforming  to  these  requirements, 
upon  which  some  valuable  cost  data  are  available,  and  which 
has  been  in  use  for  a  number  of  years,  consists  briefly  of  the 
following : 

The  method  is  applicable  to  either  room-and-pillar  or  long- 
wall  mining.  At  the  special  mine  under  consideration,  the 
output  was  600  tons  per  day  and  endless  rope  haulage  was  used 
the  empty  cars  entering  the  section  at  one  end  and  the  loads 
passing  out  at  the  other,  the  empty  cars  being  taken  off  along 
the  rope  and  the  loaded  ones  attached.  The  cars  can  pass 
around  curves  with  a  relatively  small  radius  and  they  are 
automatically  detached  at  any  point  that  may  be  desired.  The 
haulage  is  a  side  rope  system,  the  rope  being  on  the  side  farthest 
away  from  the  working  places. 

When  first  started  the  miners  were  required  to  push  their 
cars  to  the  rope  and  attach  them,  but  it  was  found  that  this 
led  to  an  unequal  distribution  of  the  cars,  the  men  at  the 
beginning  of  the  section  taking  the  most  of  the  cars  and  finish- 
ing their  day's  work  first.  To  obviate  this  six  boys  were  put 
on  pushing  the  cars  whose  duty  it  was  to  see  that  the  cars 
were  equally  distributed. 

The  following  are  the  particulars  of  a  typical  installation : 

Number  of  miners  in  section 98 

Number  of  tons  per  man 6 

Number  of  men  per  place 2 

Cars  per  day 860 

Tons  per  day 600 

Weight  per  car 14  to  13  cwt. 

Tare  of  car 4|  cwt. 

Size  of  car 4X2X3  ft. 

Rails  handled  in  12-ft.  lengths weight  24  Ib.  per  yd. 


HAULAGE  COSTS 


309 


Gauge  of  road 24  in. 

Rope,  plow  steel.. . .  weight  f  Ib.  per  ft.  and  2£  in.  in  circumference 

Height  of  coal 72  in. 

Nature  of  roof Good,  sandstone 

Nature  of  section,  flat,  little  water,  good  roof  and  floor,  coal  easily 

mined,  roof  weight  helping  considerably. 
Grade,  practically  horizontal,  about  one-half  of  1  per  cent. 


Labor  in  Section  per  Day 

Deputy,  at  $3,  for  one-third  time $1 . 00 

Haulage  man,  at  $2.50 2.50 

Six  drawers,  at  $1 . 75 10 . 50 

Two  roadmen,  at  $2,  for  one-half  time 2 . 00 

One  boy,  at  $1 .50 1 . 50 

One  night  roadman,  at  $2 2 . 00 

Shifting  and  haulage  cost,  about  $210  a  month,  per  day. ...  3 . 50 

Total $23.00 

Of  these  only  part  time  of  the  roadmen  and  very  little 
(about  one-third)  of  the  deputy's  time  are  charged  against 
the  haulage,  which  gives  a  cost  on  the  tonnage  named  of  3.83c. 
per  ton.  The  distance  traveled  is  2.08  miles,  which  works  out 
at  a  rate  of  1.84c.  per  ton-mile,  which  for  a  face  haulage  com- 
pares favorably  with  the  larger  and  more  permanent  of  end- 
less-rope haulages. 

Cost  of  wire  rope. — List  prices  of  crucible  cast  steel  rope  of 
either  standard  or  lang  lay  were  quoted  in  1920  as  follows : 


Price 
per 
Foot 

Diameter 
in 
Inches 

Approximate 
Weight 
per  Foot 

Approximate 
Strength 
in  Tons 

Working 
Load 
in  Tons 

Diameter 
of  Drum 
in  Feet 

$0.60 

li 

3  55 

63 

12.6 

11 

0.51 

H 

3 

53 

10.6 

10 

0.43 

U 

2.45 

46 

9.2 

9 

0.36 

li 

2 

37 

7.4 

8 

0.29 

1 

1.58 

31 

6.2 

7 

.22^ 

i 

1.20 

24 

4.8 

6 

.17 

4~ 

.89 

18.6 

3.7 

5 

.141 

H 

.75 

15.4 

3.1 

4f 

.12 

1 

.62 

13 

2.6 

41 

.10 

T6 

.50 

10 

2 

4 

.08 

i 

.39 

7.7 

1.5 

31 

.06^ 

A 

.30 

5.5 

1.1 

3 

.05^ 

! 

.22 

4.6 

.92 

2f 

.04? 

T6 

.15 

3.5 

.70 

2* 

.04 

A 

.125 

2.5 

.50 

If 

310 


COAL  MINING  COSTS 


List  prices  of  plow  steel  scale  lay  rope  were  quoted  as 
follows : 


Price 
per 
Foot 

Diameter 
in 
Inches 

Approximate 
Weight 
per  Foot 

Approximate 
Strength 
in  Tons 

Working 
Load 
in  Tons 

Diameter 
of  Drum 
in  Feet 

$1.30 

If 

4.85 

112 

22 

7 

1.08 

If 

4.15 

94 

19 

6.5 

.93 

i| 

3.55 

82 

16 

6 

.79 

H 

3 

72 

14 

5.5 

.65 

ij 

2.45 

58 

12 

5 

.54 

if 

2 

47 

9.4 

4.5 

.43 

l 

1.58 

38 

7.6 

4 

.34 

1 

1.20 

29 

5.8 

3.5 

.26 

I 

.89 

23 

4.6 

3 

.19 

I 

.62 

15.5 

3.1 

2.5 

.16 

A 

.50 

12.3 

2.4 

2.25 

.14 

1 

.39 

10 

2 

2 

.13 

A 

.30 

8 

1.6 

1.75 

.121 

1 

.22 

5.75 

1.15 

1.50 

•  12| 

A 

.15 

3.8 

.76 

1.25 

.12 

1 

.10 

2.65 

.53 

1 

Wire  rope  lubrication. — Wire  rope  deteriorates  with  use, 
but  not  with,  age  when  properly  cared  for;  but  the  rate  of 
deterioration  depends  in  large  measure  upon  the  character  of 
the  metal  used,  the  construction  of  the  rope,  the  diameter  of 
the  drums,  sheaves  and  pulleys  over  which  it  operates,  and 
to  a  still  greater  degree  upon  how  it  is  lubricated.  A  rope  may 
be  made  with  great  accuracy  and  meet  every  specification  that 
human  ingenuity  can  devise,  but  if  not  properly  protected  from 
the  elements  which  may  attack  its  constituent  parts,  value  and 
the  desired  economy  cannot  be  secured.  Consequently,  the 
protection  and  lubrication  of  wire  ropes  are  of  much  impor- 
tance. This  applies  to  cables  lying  idle  as  well  as  to  those  in 
service,  since  rust  destroys  as  effectively  as  hard  work. 

The  question  of  efficient  lubrication  has  recently  assumed 
a  position  of  considerable  importance.  The  manufacturer  may 
use  the  best  technical  knowledge  at  his  disposal  and  be  most 
careful  in  selecting  the  material  which  goes  into  his  product, 
but  he  cannot  be  expected  to  estimate  beforehand  the  rapid 


HAULAGE  COSTS  311 

and  varying  degrees  of  deterioration  of  the  rope  when  it  is 
in  the  hands  of  the  user,  who  it  often  happens  does  not  fully 
appreciate  that  conservation  of  the  life  of  the  cable  is  entirely 
under  his  direction.  Even  the  most  perfect  rope  can  be  used 
under  such  severe  conditions  and  with  such  lack  of  attention 
that  it  will  have  a  short  life,  while  one  of  much  inferior  quality, 
used  under  the  same  conditions  of  service  but  carefully  taken 
care  of  in  the  way  of  lubrication,  will  outlive  the  higher  grade 
rope. 

A  careful  investigation  of  all  steel  cables  used  ir  mine 
service  has  shown  that  no  two  manufacturers  are  using  the 
same  grade  of  lubricant,  and  an  analysis  of  the  materials  most 
commonly  employed  and  recommended  for  this  class  of  work 
shows  that  tar,  graphite  and  other  fillers  are  used  in  large 
proportions,  this  no  doubt  for  the  purpose  of  developing  heavy, 
adhesive  mixtures,  which  are  commonly  sold  under  the  names 
of  ''rope  shield,"  "rope  dressing"  and  "protectors."  Such 
materials  have  the  effect  of  only  partially  protecting  the 
external  parts  of  the  rope,  and  at  low  temperatures  will  crack 
and  peel.  This  may  be  noticeable  only  in  spots,  but  it  has 
the  effect  of  permitting  the  deteriorating  elements  to  attack 
the  internal  portions  of  the  cable,  and  instances  are  not  infre- 
quent of  ropes  suddenly  breaking  while  the  visible  wires  show 
no  signs  of  deterioration.  It  is  invariably  found  upon  examina- 
tion in  such  cases  that  the  internal  wires  have  perished  by 
corrosion. 

In  some  instances  ordinary  black  oil,  or  what  is  commonly 
known  as  "waste  oil,"  is  used.  Both  of  these  materials  are 
worthless  as  a  rope  lubricant,  as  neither  will  cling  to  the  outer 
surfaces,  penetrate  to  the  core,  nor  resist  the  effects  of  moisture 
or  other  damaging  elements.  Where  such  materials  are  used, 
frequent  applications  are  necessary ;  and  owing  to  their  charac- 
teristics, they  afford  no  lubrication  to  sheaves,  drums  and 
pulleys,  and  are  either  thrown  off  or  evaporated  within  a  few 
hours  after  being  applied,  thus  leaving  the  rope  at  the  mercy 
of  the  elements  and  of  frictional  wear. 

Many  high-grade  and  expensive  greases  have  been  used 
which  perhaps  by  laboratory  analysis  are  shown  to  possess  the 
qualifications  necessary  for  resisting  the  effects  of  moisture 
and  chemicals,  but  which  on  account  of  the  high  speed  at  which 


312  COAL  MINING  COSTS 

ropes  are  often  run  and  the  consequent  stress  and  vibration  to 
which  they  are  subjected  are  readily  thrown  off  and  at  no 
period  after  application  afford  more  than  a  temporary  pro- 
tection to  the  external  surfaces  of  the  rope. 

It  rarely  occurs  nowadays  that  the  full  efficiency  of  a  rope 
is  developed,  owing  to  service  conditions  requiring  it  to  come 
into  contact  with  water  containing  deteriorating  elements 
which  have  the  effect  of  quickly  producing  corrosion  and  con- 
sequent brittleness  of  the  wires.  Furthermore,  the  higher  the 
carbon  content  of  the  metal  the  more  susceptible  are  the  wires 
to  this  corrosion. 

A  series  of  practical  tests  conducted  by  a  number  of  high- 
grade  technical  men  has  demonstrated  that  the  use  of  a  pooi 
or  unsuitable  lubricant  will  often  do  more  to  lessen  the  dura- 
bility of  a  rope  than  using  no  lubricant  at  all  and  that  the  cost, 
of  a  lubricant  that  will  meet  all  operating  conditions  is  trifling 
as  compared  with  the  saving  which  its  use  makes  possible. 

An  efficient  wire  rope  lubricant  must  be  free  from  any 
material  that  will  attack  the  constituent  parts  of  the  rope, 
must  remain  soft  and  pliable  under  all  atmospheric  conditions 
and  must  not  be  subject  to  evaporation.  It  must  be  insoluble 
in  water,  so  as  to  prevent  water  from  coming  into  direct  con- 
tact with  the  surfaces  of  the  rope,  and  must  be  unaffected  by 
water  heavily  charged  with  the  chemicals  that  are  encountered 
in  mine  operation.  It  must  be  of  such  a  nature  that  it  will 
penetrate  between  the  wires  in  the  strands  and  between  the 
strands  to  the  core  of  the  rope  preserving  the  latter  as  well 
as  the  metal  which  surrounds  it.  It  must  not  be  subject  to 
decomposition  under  the  severest  conditions  of  wear  and  must 
possess  great  adhesiveness  so  that  it  cannot  be  thrown  off  by 
any  force.  It  must  be  a  material  that  will  not  harden  or  peeJ 
either  through  too  frequent  application  as  is  often  the  case  or 
under  low  temperatures,  and  must  be  of  such  a  nature  as  to 
permit  of  it  being  liquefied  to  a  consistency  to  permit  of  appli- 
cation being  made  while  quite  thin. 

In  addition  to  meeting  these  requirements,  the  lubricant 
used  for  wire  ropes  should  possess  characteristics  which  permit 
of  its  showing  equal  efficiency  as  a  lubricant  for  sheave  wheels, 
drums,  pulleys  or  other  machine  elements  over  which  ropes  are 
liable  to  pass.  And  to  secure  the  greatest  efficiency  this  same 
material  should  be  used  to  lubricate  each  wire  in  the  strand 


HAULAGE  COSTS  313 

and  to  saturate  the  core  during  the  process  of  manufacture, 
as  it  often  occurs  that  with  the  use  of  one  material  in  the 
manufacture  and  another  after  the  rope  has  been  put  in  service, 
the  two  being  of  decidedly  different  nature,  no  lubrication  can 
be  effected  on  account  of  one  material  preventing  the  other 
from  adhering  or  penetrating  to  the  interior  of  the  rope. 

Too  little  attention  has  been  given  to  the  initial  lubrication 
of  wire  rope,  particularly  as  regards  proper  saturation  of  the 
hemp  core.  It  has  been  the  general  practice  to  pass  the  hemp 
center  through  the  lubricant  used  at  the  time  the  rope  is  laid 
up.  When  applied  in  this  manner,  the  greater  part  of  the 
lubricant  is  forced  out  between  the  strands  and,  in  the  case 
of  some  grades  of  lubricant,  drips  entirely  away  from  the 
external  surfaces.  The  object  of  inserting  the  hemp  center  is 
to  increase  the  flexibility  of  the  rope,  and  the  deterioration  of 
this  element  has  the  same  effect  as  the  deterioration  of  the 
material  surrounding  it.  If  the  hemp  center  is  thoroughly 
lubricated  before  being  inserted  into  the  rope,  it  will  act  as 
a  container  of  the  lubricant  and  will  assist  in  distributing  it 
to  all  parts  of  the  cable ;  also  if  the  proper  lubricant  is  period- 
ically applied  it  will  penetrate  and  maintain  the  hemp  center 
as  a  continuous  lubricator. 

Another  important  feature  in  connection  with  wire  rope 
lubrication  which  does  not  generally  receive  the  proper  atten- 
tion is  the  method  of  application.  In  some  instances  the 
lubricant  is  poured  onto  the  rope  either  at  the  sheave  wheel 
or  at  the  drum;  in  other  instances  it  is  applied  with  a  brush. 
Both  of  these  methods  are  crude  and  wasteful. 

The  proper  way  to  apply  the  lubricant  is  to  use  a  split  box 
large  enough  to  hold  about  25  Ib.  of  the  lubricant  to  be  used. 
This  should  be  constructed  with  a  hole  in  the  center  large 
enough  for  the  passage  of  the  largest  rope  in  the  mine,  and 
when  coating  smaller  cables  an  old  rubber  pump  valve  or  a 
piece  of  ordinary  burlap  wrapped  around  the  rope  may  be 
employed  to  act  as  a  wiper  and  regulate  the  thickness  of  the 
application.  These  boxes  can  be  constructed  so  as  to  be  used 
on  either  horizontal  or  vertical  ropes. 

The  lubricant  should  be  thoroughly  liquefied  in  a  metal 
container,  and  after  it  is  poured  into  the  split  box  the  rope 
should  be  permitted  to  pass  slowly  through  it.  In  this  manner 
a  uniform  and  economical  application  can  always  be  made. 


314  COAL  MINING  COSTS 

When  possible  the  external  surface  of  the  rope  should  be  dry 
when  the  application  is  made. 

Comparative  costs  of  different  systems  of  haulage. — Many 
of  the  opinions  expressed  in  regard  to  relative  economy  of  the 
different  systems  of  haulage  are  founded  largely  on  prejudice, 
with  little  or  no  basis  for  accurate  comparison.  The  relative 
costs  of  haulage  by  all  systems  depend  upon  their  intelligent 
installation  and  handling,  and  with  equal  energy  and  experi- 
ence the  difference  in  the  cost  of  haulage  by  any  system  will 
often  be  only  a  fraction  of  a  cent.  If  the  animal  haul  can  be 
kept  short  and  the  grades  not  too  steep,  the  mule  or  horse  in 
gathering  service  is  a  close  competitor  with  the  locomotive 
except  where  the  three-  or  four-ton  mine  car  enables  the  loco- 
motive to  get  more  coal  each  time  it  makes  a  trip  to  the  room. 

An  important  factor  in  securing  maximum  economy  in 
haulage  lies  in  correctly  delimiting  the  line  of  secondary  haul- 
age, or  as  it  is  more  commonly  known,  the  gathering.  This 
will  vary  somewhat  according  to  local  conditions  at  the  dif- 
ferent mines  but  a  good  maximum  to  set  between  the  working 
face  and  the  main  haulage  switch  is  1000  ft. 

Where  long  distances  have  to  be  covered  on  secondary 
haulage  motors  are  substantially  the  most  economical.  One 
instance  is  on  record  of  a  6-ton  motor  on  secondary  haulage 
work  handling  60  cars  per  shift  over  grades  ranging  from  5 
to  17  per  cent  and  distributing  to  14  different  entries,  none  oi 
which  were  in,  over  500  ft. 

Where  approximate  values  are  required  for  general  pre- 
liminary estimates,  the  following  figures  were  those  commonly 
used  by  the  German  engineers  about  1912 :  Working  costs  for 
continuous  current  locomotives  with  overhead  contact  lines, 
0.9  to  1.2c.  per  ton-mile;  for  single-phase  locomotives  with 
overhead  contact  line,  0.9  to  1.2c.  per  ton-mile ;  for  accumulator 
locomotives,  1.8  to  2.1c.  per  ton-mile. 

The  working  costs  are  here  based  upon  the  output  in  ton- 
miles;  i.e.,  the  costs  stated  are  those  incurred  in  hauling  one 
ton  of  material  over  a  track  one  mile  in  length.  The  above 
figures,  which  can  be  reduced  under  favorable  conditions, 
enable  any  expert  acquainted  with  his  own  working  costs  to 
ascertain  by  comparison  whether  he  would  be  able  to  effect 
economy  in  his  own  installation  by  the  introduction  of  electric 
haulage. 


HAULAGE  COSTS 


315 


Comparison  of  all  systems. — One  of  the  most  exhaustive 
comparative  cost  studies  of  all  types  of  haulage  (except 
animal)  that  has  come  to  the  attention  of  the  author  was  that 
appearing  in  a  German  technical  journal  of  a  number  of  years 
ago,  the  results  of  which  are  given  herewith  together  with 
a  chart,  Fig.  21,  showing  graphically  the  percentage  of  power, 
wages,  interest  and  depreciation,  repairs  and  supplies  for  each 
of  the  different  systems. 


1   1. 


Interest  & 
Depreciation 


f^wer 


GRAPHIC  COSf  CHART 

FIG.  21. — Distribution  of  costs  for  all  kinds  of  haulage  except  animal. 

ELECTRIC  LOCOMOTIVE 

Conditions—Tonnage,  2530  per  day  of  16  hr.;   distance  3f  mi.;   speed, 
7|  mi.  per  hr.     Material  said  to  be  ore. 

Cost  of  Plant — Four  locomotives,  including  transformer  apparatus  a.nd 
trolley  wire,  $54,748. 

Working  Expenses — $25,911,  distributed  as  follows: 

Approximate 
Per  Cent  of  Total 

Interest  and  depreciation,  10  per  cent $5474  21 

Upkeep  of  locomotives 2623  10 

Upkeep  of  trolley  wires,  etc 1810  7 

Wages 5975  23 

Power 8399  33 

Oil  and  Waste 1630  6 

The  operation  cost  worked  out  at  %c.  per  ton-mile. 


316  COAL  MINING  COSTS 

STORAGE  BATTERY 

Conditions — Tonnage,  2400  per  day  of  16  hr.;  distance,  1200  yd.;  speed, 
6|  mi.  per  hr. 

Cost  of  Plant — Cost  of  five  locomotives  (20-hp.),  weighing  6|  tons, 

also  of  transformers,  switchboard  and  reserve $14,600 

Accumulators 10,219 

Cable,  rooms  for  transformers  and  charging 7,542 

Total $32,361 

Working  Expenses — Annual  amount,  $14,133. 

Approximate 
Per  Cent  of  Total 

Depreciation. $1460  10 

Parts  for  batteries 1469  lOf 

Interest  on  total  first  cost 1095  8 

Wages  of  drivers 3581  25| 

Brakemen's  wages 992  7 

Attendance  on  transformers  and  batteries.. .     2671  19 

Upkeep  and  cleaning 569  4 

Oil  and  waste 141  1 

Power,  186,349  kw.-hr 1790  12| 

Acid,  distilled  water,  etc 355  2| 

The  operation  costs  were  3c.  per  ton-mile,  but  it  is  con- 
sidered that  the  conditions  were  not  favorable. 


BENZOL  (GASOLINE)  LOCOMOTIVE  PLANT 

Conditions — Tonnage,  1250  per  day  of  16  hr.;  distance,  1000  yd.,  sloping 
toward  shaft  at  4  per  cent  grade;  speed  of  3|  to  5|  mi.  per  hr. 

Cost  of  plant $8930 

Four  locomotives,  8-hp $6813 

Filling  apparatus 73 

Engine  shed  for  six  machines 2044 

Working  expenses — for  six  months,  $587. 

Approximate 
Per  Cent  of  Total 

Interest  and  depreciation $87  15 

Upkeep  of  locomotives 180  31 

Wages  of  engine  drivers .- 136  23 

Wages  of  brakemen,  etc 68  11 

Benzol 97  17 

Oil  and  waste 19  3 

This  operation  works  out  at  3c.  per  ton-mile. 


HAULAGE  COSTS  317 


COMPRESSED-AlR   HAULAGE 

Conditions — Tonnage,  1242  tons  in  16  hr.;  distance  1|  nil. 
Cost  of  Plant — Four  12-hp.  locomotives  weighing  5|  tons,  hauling  40  to 
50  cars  of  17  cwt.  gross,  compressor,  plant  and  accessories,  $14,500. 
Working  Expenses — Annual  amount,  $9391. 

Approximate 
Per  Cent  of  Total 

Interest  and  depreciation  at  10  per  cent $1460  16 

Repairs,  upkeep,  oil  and  waste 608  6 

Wages,  brakemen  and  switchmen 851  9 

Wages  of  drivers  (5) 1971  21 

Cost  of  attendance  on  plant 365  4 

Consumption  of  power  (170  hp.,  16  hr.  per 

day,  300  days) 4136  44 

This  works  out  at  2^c.  per  ton-mile. 


ROPE  HAULAGE 

Conditions — Tonnage,  1690  in  18  hr.;  distance,  1400  yd.,  main  road 
served  by  a  number  of  branches.  Gradient  of  5  per  cent  for  700  yd.,  including 
right-angled  turn. 

Cost  of  Plant — $5092. 

Total  cost  of  installing $2929 

Rope 973 

Share  of  outlay  on  haulage  engine  and  buildings 2190 

Working  Expenses — Annual  amount,  $7526. 

Approximate 
Per  Cent  of  Total 

Depreciation  and  interest $496  6£ 

Wear  and  tear  on  rope 486  6| 

Wages  of  hookers-on 3504  46 

Two  engine  drivers 734  10 

Roadman 365  5 

Upkeep,  repairs,  oil,  cleaning,  waste 116  2 

Cost  of  power 1825  24 

This  cost  works  out  at  l%c.  per  ton-mile. 


318  COAL  MINING  COSTS 


SECOND  ROPE  HAULAGE 

Conditions — Tonnage,  2100  in  10  hr.;  distance,  3  mi.  on  a  level  track. 

Cost  of  Plant — Including  engine,  250-hp.,  and  accessories,  $48,666. 

Working  Expenses — Annual  amount,  $15,069. 

Approximate 
Per  Cent  of  Total 
Interest  and  depreciation  at  10  per  cent. . . .  $4866  32 

Wear  and  tear  of  rope 851  5£ 

Upkeep,  repairs,  cleaning,  oil  waste 618  4 

New  parts  for  rope  haulage  and  rope  clips ..       837  5^ 

Wages  of  hookers-on 4088  27 

Engine  driver 296  2 

Cost  of  power 3513  23 

This  works  out  to  %c.  per  ton-mile. 

OVERHEAD  CHAIN 

Conditions — Tonnage,  1750  per  day  of  16  hr.;  level  track  1£  mi.  long. 
Cost  of  Plant — Machinery,  including  65-hp.  engine,  $6082;   chain,  $2929; 
total,  $9011. 

Working  Expenses — Annual  amount,  $13,828. 

Approximate 
Per  Cent  of  Total 

Interest  and  depreciation $876  6 

Interest  and  depreciation  of  chain 525  4 

Wages  of  hookers-on 7878  57 

Repairs  and  upkeep  of  plant Ill  1 

Oil  and  waste 141  1 

Cost  of  power 1786  13 

Upkeep  of  haulage  road 2501  18 

This  is  almost  exactly  2c.  per  ton-mile. 

Animal,  compressed-air  and  electric  haulage  costs. — A  most 
interesting  and  valuable  comparison  of  the  cost  of  animal,  com- 
pressed air  and  electric  haulage  at  a  mine  in  Western  Pennsyl- 
vania, producing  35,000  tons  a  month  was  worked  up  in  1912. 
The  grades  in  this  mine  are  variable,  some  of  them  in  favor 
of,  and  some  of  them  against  the  loads,  and  are  in  places  as 
steep  as  6  per  cent.  The  capacity  of  the  mine  cars  is  two 
tons,  and  their  empty  weight  2700  Ib.  In  the  case  of  the  elec- 
trically operated  mine,  the  car  capacity  was  not  definitely 
stated,  but  presumably  it  was  about  3000  Ib.,  and  the  weight 
of  the  car  itself  about  1300  Ib.  A  comparison  of  figures  for 
the  two  cases  is  given  in  Tables  I  and  II. 


HAULAGE  COSTS 


319 


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Total  labor  

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Repairs  and  maintenance  
Power  
Interest  6%  on  $10,000  for  one  month.  .  . 
Interest  6%  on  $5900  for  one  month 
Depreciation  eight  years  
Depreciation  10%  

320  COAL  MINING  COSTS 

TABLE  II 
RELATIVE  COST  PER  MILE  AND  DATA  FOR  ESTIMATE 

Average  haul  with  electric  locomotive 3600  ft. 

Average  haul  with  compressed  air  locomotive 2100  ft. 

Total  ton-miles,  electric  locomotive 2758 

Total  ton-miles,  compressed  air  locomotive 4138 

Total  cost  per  ton-mile,  electric  locomotive 14 .  68c. 

Total  cost  per  ton-mile,  compressed-air  locomotive. . . .  14. 57c. 

Working  two  shifts  is  a  decided  advantage  to  either  electric 
locomotives  or  compressed-air  locomotives  as  compared  with 
mules,  because  mules  cannot  be  worked  two  shifts,  whereas 
locomotives  can,  thereby  distributing  the  increased  interest  and 
depreciation  charges  over  double  the  number  of  hours.  As  a 
matter  of  fact,  the  compressed-air  locomotives  at  this  mine 
only  worked  one  shift,  but  the  costs  per  hour  and  per  ton 
have  been  held  exactly  the  same  as  they  actually  were  with  the 
exception  that  the  interest  and  depreciation  are  distributed 
over  553  hours  per  month  and  over  the  amount  of  coal  that 
would  have  been  gathered  in  553  hours  had  the  locomotives 
continued  to  work  at  the  same  rate  which  they  maintained  for 
nine  hours  per  day  and  six  days  in  the  week.  This  was  done 
in  order  to  make  the  costs  more  truly  comparable. 

It  should  be  observed  in  connection  with  the  above  figures 
for  cost  per  ton  and  ton-mile  that  neither  of  these  costs  is  in 
general  an  accurate  basis  for  testing  the  comparative  merits 
of  the  two  types  of  haulage,  particularly  as  regards  gathering 
service,  although  the  same  considerations  apply  to  a  more 
limited  extent  in  connection  with  main  haulage.  When  gath- 
ering, the  locomotive  necessarily  spends  most  of  its  time  in 
assembling  the  trip,  while  the  ton  mileage  is  run  up  by  the  long 
haul  between  the  point  where  the  coal  is  gathered  and  the 
point  where  it  is  delivered.  The  expense  of  haulage  does  not 
increase  directly  in  proportion  to  the  length  of  the  haul  be- 
cause the  time  spent  in  gathering  remains  a  constant  regard- 
less of  the  length  of  the  direct  run.  Usually  if  an  analysis  is 
made  of  the  power  expended  and  time  spent  in  gathering  and 
in  hauling  the  gathered  cars  to  the  terminus,  it  will  be  found 
that  the  strictly  gathering  service  is  the  most  expensive  part 
of  the  work. 


HAULAGE  COSTS 


321 


Changing  those  items  in  the  foregoing  table  which  are  either 
manifestly  not  chargeable  to  the  gathering  type  of  locomotive 
or  else  need  correction,  Table  IV  is  derived.  The  figures  are 
for  the  same  locomotives  working  under  the  same  conditions 
as  before,  but  with  the  engineers  and  brakemen  receiving  the 
same  rate  of  pay,  with  the  charge  for  power  increased  from 
$20  to  $50  per  month,  and  the  item  depreciation  8  yr.,  $10,000 
raised  to  $84.20  per  month.  The  reason  for  changing  the  rate 
of  $20  per  month  for  power  is  explained  later. 

Based  on  the  corrected  and  equalizing  figures  of  Table  IV 
the  cost  per  net  ton-mile  would  be  17.55c.  for  electric  locomo- 
tive and  12.32c.  for  the  compressed-air  locomotive. 

In  the  same  mine  where  the  compressed-air  gathering  loco- 
motives operate,  horses  are  used  for  gathering  coal  in  another 
part  of  the  mine.  The  coal  gathered  by  both  the  small  air 
locomotives  and  the  horses  is  hauled  to  the  foot  of  the  shaft 
by  large  compressed-air  locomotives.  The  cost  of  gathering  by 
horses  in  this  mine  is  as  follows : 


TABLE  III 
COSTS  OP  GATHERING  BY  HORSES 


Day 

Week 

Cost 
per  Ton 

Nine  drivers  at  $2  .  60  per  day  

$23.40 

$140  40 

One  driver  at  $2  70  per  day  

2  70 

16  20 

Total  labor 

$156  60 

$0  0318 

Stable  expense  

83  60 

0  0170 

Thirteen  horses,  $3250  depreciation  5  years  .  . 
Interest,  6%  on  $3250  



12.50 
3.75 

0.0025 
0  0007 

$256.45 

$0.0520 

With  the  ten  drivers,  eleven  work  horses  and  two  spares, 
4918  tons  of  coal  were  gathered  per  week  and  hauled  an  aver- 
age distance  of  900  ft.,  giving  the  cost  per  ton  as  stated  above 
in  detail  and  a  total  cost  per  ton-mile  of  30.6c. 

The  cost  per  ton  for  gathering  with  mules  or  horses  in  the 
electrically  operated  mine  was  as  follows : 


22                                     COAL  MINING  COSTS 

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HAULAGE  COSTS  323. 

TABLE  V 
COSTS  OF  MULE  HAULAGE 

Average  distance  of  mule  haul 1200  ft. 

Drivers,  250  hr.  at  30c $75 . 00 

Drivers,  398  hr.  at  18c 71 . 64 

Drivers,  294  hr.  at  19c 55.86  Cost  per 

Drivers,  310  hr.  at  20c 62 . 00  Ton 

$264.50  $0.0646 

Stable  boss 35.00      0.0086 

Blacksmith,  shoeing 30 . 00      0 . 0074 

Feed 79.20      0.0193 

Miscellaneous  supplies 18 . 52      0 . 0043 

Depreciat  on,  5  years 25 . 00      0 . 0062 

Interest,  6%  on  $1500 7. 50      0. 0018 


Total $459.72     $0. 1122 

Total  cost  per  ton  mile 49 .  4c. 

The  coal  gathered  by  the  mules  in  this  mine  was  hauled  to 
the  pit  mouth,  an  average  distance  of  2500  ft.,  by  an  electric 
locomotive. 

The  comparative  figures  for  main  haulage  in  the  two  mines 
under  consideration,  one  using  electricity  and  the  other  com- 
pressed air  are  given  in  Table  VI. 

The  electric  locomotive  moved  the  coal  an  average  distance 
of  2500  ft.  and  the  compressed-air  locomotive  hauled  it  for  au 
average  distance  of  3400  ft.,  so  the  costs  per  ton-mile,  are  as 
follows : 

For  electricity: =  1938  ton-miles, 


1939  =10.83c.  per  ton  mile. 

3400X19,688 

For  air :  =  12,677  ton-miles, 

5280 

386  31 
12677  =  3.05c.  per  ton  mile. 

In  explanation  of  the  correction  of  the  figure  of  $20  per 
month  for  power  as  given  in  Table  I  and  increased  to  $50  in 
Table  IV,  it  should  be  noted  that  the  costs  of  power  per  ton-mile 


324 


COAL  MINING  COSTS 


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HAULAGE  COSTS  325 

for  gathering  and  main  haulage  as  previously  considered  were 
as  follows : 

Cost  for  power  in  gathering $20.00 

Ton  mileage  as  calculated 2758 

$20-00  =  $0 . 00726  per  ton-mile. 

2758 

And  for  main  haulage: 

con  00 

^P  =  $0 . 01032  per  ton-mile. 
1939 

It  will  be  seen  from  these  figures  that  the  cost  per  ton-mile 
is  42  per  cent  greater  for  the  main  haulage  than  for  the  gather- 
ing service. 

The  costs  of  power  per  ton-mile  for  the  same  two  classes 
of  service  at  the  compressed-air  operated  mine  were : 

For  gathering :  '      =  $0 . 0284  per  ton-mile. 

41oo 

$80  18 

For  main  haulage :    :        —  =  $0 . 0063  per  ton-mile. 
12677 

These  figures  show  the  cost  for  gathering  with  compressed- 
air  locomotives  to  be  4.48  times  as  great  as  for  main  haulage. 
There  are  many  reasons  why  the  cost  for  power  per  ton-mile 
for  gathering  should  be  much  more  than  for  main  haulage,  but 
none  why  it  should  be  less,  unless  under  such  a  condition  as 
where  all  the  main  haulage  grades  are  against  the  loads  and 
all  the  gathering  is  down  hill. 

In  all  the  above  figures  it  is  the  ton  of  coal  moved  one 
mile  in  the  desired  direction  that  is  considered.  In  actual 
service,  the  cars  and  the  locomotive  must  be  moved  as  well 
as  the  coal,  and  because  the  car  and  the  locomotive  must  move 
approximately  twice  as  far  as  the  coal  in  order  to  get  the 
empty  cars  back  to  the  loading  place,  the  gross  ton-mileage  is 
nearly  twice  the  net  ton  mileage  for  main  haulage.  For  gath- 
ering work  the  gross  ton-mileage  will  be  from  four  to  six 
times  the  net  ton-mileage  because  in  gathering  cars  from  the 
rooms  the  locomotive  must  make  many  movements  with  only 
one  car  or  without  any  cars.  While  doing  this  work  the  per- 
centage of  net  to  gross  tonnage  is  exceedingly  low  because  the 


326  COAL  MINING  COSTS 

locomotive  itself  forms  a  large  part  of  the  total  weight  of  the 
moving  train. 

Any  remaining  inconsistency  in  the  relative  costs  of  main 
haulage  and  gathering  service  with  the  compressed-air  loco- 
motives is  readily  accounted  for  by  considering  the  following 
adverse  conditions  of  gathering  as  compared  with  main  haul- 
age: Ordinarily  the  gathering  locomotives  has  poorer  track 
than  the  main-haulage  locomotive,  a  greater  percentage  of 
curves,  more  frequent  stops  and  starts,  and  a  somewhat  lower 
efficiency,  because  of  handling  only  one  car  at  a  time  instead 
of  being  loaded  approximately  up  to  its  capacity. 

From  the  foregoing  considerations  it  seems  certain  that  the 
figure  of  $20  as  the  cost  of  power  for  the  electric  gathering 
locomotive  needs  correction;  more  especially  as  this  locomotive 
operated  two  shifts  per  day,  while  the  main-haulage  locomotive 
operated  only  one  shift. 

The  conclusion  that,  as  it  cost  only  lOc.  per  ton  to  gather 
and  haul  4045  tons  of  coal  per  month  with  one  locomotive 
working  two  shifts,  and  as  it  cost  16.34c.  per  ton  to  gather 
4095  tons  of  coal  with  mules  and  haul  it  2500  ft.  to  the  pit 
mouth,  there  is,  therefore,  an  advantage  of  not  less  than  4c. 
per  ton  in  gathering  by  electric  locomotives,  does  not  seem  to 
have  sufficient  support  to  make  it  generally  true. 

In  the  mine  using  compressed-air  haulage,  coal  was  gathered 
and  hauled  by  mules  an  average  distance  of  900  ft.  at  a  cost 
of  5.2c.  per  ton  and  was  afterward  hauled  an  average  distance 
of  3400  ft.  by  main-haulage  compressed-air  locomotives  at  a 
cost  of  1.97c.  per  ton,  giving  a  total  cost  for  gathering  and 
hauling  of  7.17c.  per  ton,  a  lower  figure  than  that  in  the  electric 
haulage  mine.  This  figure  is  also  slightly  lower  than  the  results 
achieved  with  compressed-air  gathering  and  main-haulage  loco- 
motives in  sections  of  the  same  mine  where  the  gathering 
locomotives  worked  but  one  shift,  as  is  shown  by  the  follow- 
ing costs: 

Gathering  by  compressed-air  locomotives 6.72c.  a  ton 

Main  haulage 1 . 97c.  a  ton 


Total  cost 8 . 69c.  a  ton 

Thus  it  is  seen  that  the  coal  was  delivered  to  the  shaft 
bottom  more  cheaply  by  mules  and  main-haulage  locomotives 


HAULAGE  COSTS  327 

than  by  gathering  locomotives  and  main-haulage  locomotives; 
but  in  order  to  achieve  these  results  with  the  mules  it  was 
necessary  to  keep  the  mule  haul  down  to  900  ft.  or  less  in  order 
to  enable  the  animals  to  gather,  as  they  did,  41.6  two-ton  cars 
per  mule.  If  the  mules  had  to  haul  the  coal  an  average  dis- 
tance of  1200  ft.,  as  did  the  gathering  locomotives,  and  had 
been  forced  to  encounter  the  adverse  grades  that  the  locomo- 
tives did,  the  cost  for  mule  haulage  would  have  been  50  per 
cent  greater,  throwing  the  balance  again  in  favor  of  the  gath- 
ering locomotive.  In  other  words,  it  was  possible  to  obtain 
the  good  results  that  were  obtained  with  mules  in  this  mine 
only  by  working  them  in  selected  places  under  the  most  favor- 
able conditions,  all  of  which  goes  to  show  that  it  is  extremely 
dangerous  to  draw  general  conclusions  in  regard  to  the  relative 
economy  of  the  various  types  of  haulage  unless  the  conditions 
are  strictly  comparable,  or  accurate  corrections  are  made  to 
compensate  for  differing  conditions. 

Single-  and  two-stage  air  motors. — The  two-stage  com- 
pressed-air motor  will  do  from  40  to  60  per  cent  more  work 
with  the  same  amount  of  air  than  the  single-stage  machine. 
The  cost  of  power  for  this  type  motor  will  thus  be  about  30 
per  cent  less  than  with  the  single-stage,  the  compressor  and 
boiler  capacity  will  be  reduced  by  about  the  same  amount 
and  the  first  cost  of  the  installation  will  be  about  15  per  cent 
less,  while  the  motors  will  travel  substantially  further  on  one 
charge. 

A  practical  test  of  the  difference  between  single-expansion 
and  two-stage  compressed-air  locomotives  was  once  made  as 
follows:  The  same  train  was  hauled  by  the  single-expansion 
and  two-stage  locomotives  the  same  distance  over  the  same 
piece  of  track,  with  the  same  operator,  giving  the  locomotives 
as  nearly  as  possible  the  same  work  to  do  under  the  same  con- 
ditions. In  all  cases  the  trains  were  started  from  a  given  point 
and  were  allowed  to  come  to  rest  as  near  as  possible  to  an- 
other given  point  without  the  use  of  brakes.  Air  consumed 
was  determined  by  the  difference  in  pressure  recorded  at  the 
beginning  and  end  of  the  trip  by  the  gauge  on  the  main 
reservoir.  Running  conditions  were  the  same  as  would  obtain 
in  the  regular  hauling  of  coal  in  the  mine. 

Trial  1  was  conducted  at  the  Susquehanna  Coal  Co.'s  No. 


328 


COAL  MINING  COSTS 


1  colliery,  Nanticoke,  Pa.,  in  the  presence  of  Mr.  McMahon, 
Chief  Engineer,  Susquehanna  Coal  Co.,  and  Mr.  C.  B.  Hodges, 
of  the  H.  K.  Porter  Co. 

TRIAL  1 
LOCOMOTIVE  DATA 


Single- 
Expansion 

Two-Stage 

Weight     

10,000  Ib 

10,600  Ib 

Cylinders  high-pressure,  diameter 

6  in 

5  in 

Cylinders,  low-pressure,  diameter  

10  in 

Cylinders,  stroke 

10  in 

10  in 

Driving  wheels,  number  and  diameter  .... 
Working  pressure,  high-pressure  cylinder  . 
Maximum  charging  pressure  

4-23  in. 
150  Ib. 
900  Ib. 

4r-23  in. 
250  Ib. 
900  !b. 

Capacity  of  main  reservoir  

41  cu.  ft. 

41  cu  ft 

Age  of  locomotive 

8  months 

1  month 

TRAIN  AND  ROAD  DATA 

Length  of  trial  run,  feet 2200 

Average  grade,  per  cent 0 . 96 

Maximum  grade,  per  cent 2 . 00 

Track  gage,  inches 42 

Train  consisted  of  loaded  coal  cars  weighing  about  9500  Ib. 
each. 

The  excessively  high  efficiency  indicated  on  lines  Nos.  4 
and  6  of  the  table  of  log  of  runs,  hauling  four  and  five  cars 
down  grade,  may  be  due  to  the  fact  that  considerably  more 
air  is  wasted  in  the  single-expansion  locomotive  when  giving 
the  train  just  a  little  assistance  when  the  grade  is  nearly  steep 
enough  to  cause  it  to  run  down  by  itself. 

Lines  Nos.  7  and  8  show  additional  trips  up  grade  with  the 
two-stage  with  five-car  trips.  The  trip  with  the  two-stage 
(line  5)  was  made  with  the  reverse  lever  "in  the  corner, " 
using  the  air  full  stroke  all  the  way.  The  superior  efficiency 
achieved  on  trips  shown  on  lines  Nos.  7  and  8  was  the  result 
of  using  the  air  as  expansively  as  possible. 

Taking  the  total  pressure  reduction  of  the  three  round  trips 
with  the  single-expansion  locomotive  and  two-stage  locomotive 
(lines  1,  2,  3,  4,  5,  and  6),  we  find  that  the  two-stage  locomotive 


HAULAGE  COSTS 


329 


used 


1245 
2225 


=  56  per  cent  of  the  air  used  by  the  single-expansion 


engine  under  exactly  the  same  conditions  of  service,  this  average 
per  cent  for  the  entire  test  showing  a  saving  of  44  per  cent  of  the 
air  used  by  the  single-expansion  machine. 

In  making  the  test  every  care  was  used  to  obtain  reliable 
results.  The  pressure  gauge  on  the  two-stage  locomotive  was 
removed  and  placed  on  the  single-expansion  locomotive  during 
the  test  of  this  locomotive,  and  then  shifted  back  to  the  two- 
stage  for  testing  it.  This  gauge  was  a  comparatively  new  one, 
and  presumably  correct,  and  if  any  error  did  exist,  it  would 
have  been  the  same  for  both  locomotives.  The  tanks  were 
exactly  of  the  same  capacity.  The  same  engineer  operated  both 
locomotives  alternately  during  the  trials. 

Trial  2  was  conducted  at  Orient  Mine  of  the  Orient  Coke 
Co.,  Orient,  Fayette  County,  Pennsylvania,  in  the  presence  of 
Mr.  Chas.  Opperman,  of  the  Orient  Coke  Co.;  Mr.  G.  E.  Hut- 
telmaier,  of  the  H.  C.  Frick  Coke  Co. ;  Mr.  C.  B.  Hodges,  of 
the  H.  K.  Porter  Co. 

TRIAL  2 
LOCOMOTIVE  DATA 


Single- 
Expansion 

Two-Stage 

Weight  . 

9600  Ib 

10  500  Ib 

Cylinders,  high-pressure,  diameter  

6  in 

51  in 

Cylinders,  low-pressure,  diameter.. 

11  in 

Cylinders,  stroke 

10  in 

10  in 

Driving  wheels,  number  and  diameter  
Working  pressure,  high-pressure  cylinder. 
Maximum  charging  pressure 

4r-23in. 
150  Ib. 
800  Ib 

4r-23in. 
250  Ib. 
800  Ib 

Storage  tanks  

1 

1 

Capacity  of  main  reservoir  

40  26  cu  ft 

40  26  cu  ft 

Age  of  locomotive 

6  months 

2  months 

TRAIN  AND  ROAD  DATA 

Length  of  trial  run,  feet 2500 

Average  grade,  per  cent 52 

Track  gage,  inches 44 

In  this  run  there  was  a  reverse  curve  in  a  chute  leading  from  one  heading 
to  a  parallel  heading.  The  train  consisted  of  four  loaded  wagons,  each  about 
7000  Ib.;  and  six  empty  wagons,  each  about  2200  Ib. 


330 


COAL  MINING  COSTS 


LOG  OP  TRIAL  RUNS 


TANK  PRESSURES 

TIME  (P.M.) 

No. 
of 
Run 

Type  of 
Locomotive 

At 
start 

At 
finish 

Amount 
of 

Start 

Finish 

Elapsed 

Drop 

1 

Single-expansion  . 

705 

265 

440 

7:23£ 

7:27^ 

.04 

?, 

Two-stage  

740 

420 

320 

8:06* 

8'12 

05  i 

3 

Two-stage  

685 

385 

300 

8:48 

8:52f 

.04| 

No.  1  run:    Very  satisfactory. 

No.  2  run:    Very  irregular;     operator    not    so    familiar   with    two-stage 
machine,  hence  decided  to  rerun. 

No,  3  run;   Much  better  and  smoother  than  second. 


DEDUCTIONS 
Calculated  by  Mr.  G.  E.  Huttelmaier,  H.  C.  Frick  Coke  Co. 


Trial  No.  1 

Trial  No.  2 

Free  air  consumed,  cubic  feet  

1,206.92 
930.16 

2,325,400.00 
581,350.00 
625.00 
7.10 
17.61 
1,926.00 
301  .  70 
17.13 

100  per  cent 
100  per  cent 

877.76 

«:£}«•« 

2,348,350.00 
426,973.00 
454.50 
5.17 
12.94 
2,676  .  00 
159.59 
12.33 

72  per  cent 

1.38  per  cent 
28  per  cent 

Drawbar  effort  \  „.                     \  .  . 

I  Trip  832.  24  / 
Total  work  performed  in  foot-pounds  
Foot-pounds  performed  per  minute  .  .      .    . 

,  /  Feet  per  minute  "1 
Average  speed  \  ,  ,.,            •>            >  .  . 

1  Miles  per  hour    / 
Average  horsepower  developed     .    ... 

Foot-pounds  work  per  cubic  foot  free  ah*  
Free  air  consumed  per  minute  

Air  per  minute  per  horsepower 

Percentage  of  air  consumed  as  compared 
with  Trial  No   1 

Amount  of  work  per  unit  of  air  compared 
to  Trial  No   1 

Saving  of  air  effected  over  Trial  No.  1  .... 

HAULAGE  COSTS 


331 


LOG  OP  RUNS — WITH  PRESSURE  REDUCTIONS  AND  PERCENTAGE  USED  BY 
TWO-STAGE  LOCOMOTIVE 


Line 

1 
2 
3 
4 
5 
6 

7 
8 

SlNGLE-E  X  PANSION 

TWO-STAGE 

Percentage  Used 
by  Two-Stage 

No. 
Cars 

Gage 
Pressure 

Pres- 
sure 
Re- 
duction 

Time 

No. 
Cars 

Gage? 
Pressure 

Pres- 
sure 
Re- 
duction 

Time 

i 
Start 

Stop 

Start 

Stop 

3 
3 
4 
4 
5 
5 

1 

i 

690 
280 
770 
760 
805 
730 

\>tal  p 
2225  It 

. 

280 
140 
180 
540 
200 
470 

ressvm 
>.  —  3  r 

410 
140 
590 
220 
605 
260 

3:15 
3:30 
2:55 

an 

1 

3 
3 
4 
4 
5 
5 

1 

5 
5 

680 
400 
570 
250 
710 
290 

^otal  p 
1245  It 
610 
735 

400 
320 
250 
175 
290 
220 

ressurt 
.  —  3  r< 
220 
360 

280 
80 
320 
75 
420 
70 

68.3 
57.1 
54.2 
34.1 
69.5 
26.9 

Up  grade 
Down  grade 
Up  grade 
Down  grade 
Up  grade 
Down  grade 

Up  grade 
Up  grade 

2:40 

2:50 

n 
1 

2252 

;  reducti 
sund  tri] 

1245 

reductic 
>und  trip 
390 
375 

3:15 

Compressed-air  and  animal  haulage  costs. — The  Consolida- 
tion Coal  Co.,  in  1902,  conducted  an  investigation  into  the 
relative  cost  of  mule  and  compressed-air  haulage,  the  results 
of  which  were  described  in  the  transactions  of  the  A.I.M.M.E., 
Vol.  34,  p.  144. 

Five  of  these  machines  were  placed  in  the  company's  Ocean 
No.  3  mine  (Hoffman),  displacing  a  number  of  mules,  but  leav- 
ing 19  still  working.  This  opportunity  was  embraced  to  make 
a  close  comparison  between  the  two  methods  of  gathering. 

The  mules  working  in  the  North  Heading  and  the  South 
Heading  deliver  their  cars  directly  to  the  rope  on  the  slope 
The  other  mule  routes  deliver  to  the  heavy  motors,  as  do  all 
the  motor  routes.  The  mules  used  weigh  from  1200  to  1400  lb., 
and  pull  an  average  of  2.4  long  tons. 

The  following  table  shows  the  work  performed  by  the  mules 
during  a  period  of  18%  working  days  in  the  month  of  Decem- 
ber, 1902: 


332 


COAL  MINING  COSTS 


Route 

Cars 
Moved 

Average 
Haul, 
Feet 

Constant 

Tons 
Moved 
1000  Feet 

South  heading            

1119 

2900 

2.4 

7788  24 

North  heading  

268 

1300 

1000 
do 

836  16 

First  cross                  .  . 

1629 

2100 

do 

8210  16 

Second  cross 

3042 

1100 

do 

8030  88 

Third  cross             

747 

400 

717  12 

This  represents  a  total  of  339  days'  work  for  one  mule.  The 
company's  accounts  show  a  cost  of  $1.15  per  day  for  each 
day  worked  by  a  mule,  including  expense  of  replacing  worn- 
out  animals.  Drivers  are  paid  $1.98,  and  there  is  one  with 
each  mule.  This  makes  a  cost  of  $3.13  per  day  for  each  day 
worked  by  a  mule.  The  cost  per  ton  hauled  1000  ft.  would . 
therefore,  be: 


339XS3.13 


=4.15c. 


25,582.56 
For  the  work  of  the  motors  during  the  same  time  we  have . 


Route 

Cars 
Moved 

Average 
Haul, 
Feet 

Constant 

Tons 
Moved 
1000  Feet. 

Tippens 

1122 

2300 

2.4 

6  193  44 

Scobies 

1073 

2050 

1000 
do 

5  279  16 

First  Klondyke  

1147 

1835 

do 

5  046  80 

Second  Klondyke. 

1032 

1800 

do 

4458  24 

Third  Klondyke 

1147 

1865 

do 

5  138  56 

Fourth  Klondyke  

114 

1992 

do 

544  92 

Total  

26  661  12 

This  work  was  done  by  the  five  small  motors  operated  by 
compressed  air,  working  a  total  of  94  days.  This  plant  is 
supplied  with  steam  by  a  battery  of  boilers,  which  also  supplies 
steam  to  the  large  pumps.  The  plant  consists  of  the  following 
items,  with  their  approximate  first  cost: 


HAULAGE  COSTS  333 

One  straight-line  Norwalk  air-compressor,  18  and 
28  compound  steam,  18£,  13J,  and  6£  three- 
stage  air,  30-in.  stroke $5,300 

5600  ft.  of  5-in.  pipe 5,600 

3100  ft.  of  2Hn.  pipe 1,700 

1000  ft.  of  U-in.  pipe 300 

2  motors,  30,000  Ib.  each 6,000 

5  motors,    8,000  Ib.  each 10,000 

Estimated  proportion  of  boilers 1,000 

Installation 4,000 

$33,900 

Allowing  $3000  per  year  for  interest  and  depreciation,  to 
be  earned  in  300  working-days,  would  justify  a  charge  of  $10 
per  day  from  this  source  against  the  entire  plant. 

This  same  compressor  also  drives  the  large  motors  men- 
tioned above,  which  weigh  30,000  Ib.  each  (60,000  Ib.  for  the 
two)  ;  the  five  small  machines  weigh  8000  Ib.  each  (40,000  Ib. 
for  the  five).  Dividing  the  general  expenses  according  to  the 
weight  would  result  in  four-tenths  being  charged  against  the 
small  motors. 

These  general  expenses  may  be  summed  up  as  follows  per 
day: 

Coal,  4  tons  at  $1 $4.00 

Fireman 2 . 00 

Mechanic  in  charge  of  compressor 2 . 50 

Interest  and  depreciation 10 . 00 


$18.50 
The  cost  of  operation  of  the  five  small  motors  would  then 


be: 


5  motormen  @  $2 . 67 $13 . 35 

5  brakemen  @  $2 . 03 10. 15 

General  expenses  0.4X$18.50 7.40 

Repairs  and  oil 3 . 00 


$33.90 

Dividing  this  among  the  five  machines  would  give  $6.78  per 
day  for  each  machine  and  the  cost  per  ton  moved  1000  ft.  would 
,  6.78X94 


334  COAL  MINING  COSTS 

In  the  matter  of  community  of  service  the  motors  show  to 
great  advantage.  A  broken  down  motor  can  usually  be  re- 
paired over  night,  while  an  injured  mule  can  only  be  replaced 
by  a  new  one  that  must  usually  be  broken  in  and  inured  to 
the  work  before  he  is  thoroughly  efficient,  entailing  loss  of 
time  and  output  in  each  case. 

Electric  motor  and  animal  haulage  costs. — An  interesting 
and  valuable  comparison  of  the  cost  of  electric  and  mule  haul- 
age at  one  of  the  mines  of  the  Peabody  Coal  Co.  was  worked 
up  in  1907.  The  introduction  of  the  electric  haulage  not  only 
resulted  in  reducing  the  cost  of  production,  but  also  made  prac- 
ticable the  development  of  more  extended  operations  and 
increased  the  output  from  1400  tons  to  a  daily  average  of  2000 
tons.  The  motors  have  pulled  as  much  as  2570  tons  in  8  hr. 

Prior  to  installing  electric  haulage  there  were  16  gathering 
mules  and  17  mules  working  in  spike  teams,  pulling  to  the 
bottom,  producing  1400  tons  of  coal  per  day.  Owing  to  the 
size  of  the  cars,  grade  and  average  haul  of  1800  ft.  to  the  bot- 
tom of  the  shaft,  the  output  had  reached  its  limit  with  mule 
haulage,  and  it  was  decided  to  install  electrical  haulage.  Two 
15-ton  traction  locomotives  with  double-end  control  were  in- 
stalled. The  locomotives  have  pulled  17  loaded  cars  up  a  2l/2 
per  cent  grade  1200  ft.  long.  These  cars  weigh  when  empty 
1950  Ib.  and  hold  on  an  average  6600  Ib.  coal,  so  the  weight 
of  the  loaded  trip  would  be  over  72  tons. 

The  power  for  operating  the  motors  in  the  mine  is  supplied 
by  a  175-kw.  generator  belted  to  a  200-hp.  high-speed  engine, 
18  X  18  in.  The  generator  also  furnishes  light  for  the  under- 
ground haulage-ways. 

Steam  to  run  the  electric  plant  is  furnished  by  a  battery  of 
four  150-hp.  tubular  boilers,  which  also  furnishes  steam  for 
the  large  hoisting  engines;  but  in  order  to  make  the  proper 
comparisons  between  mule  and  electric  haulage,  the  cost  of 
two  complete  power  units  has  been  added  to  the  electrical 
equipment.  The  machinery  making  up  the  electrical  installa- 
tion is  as  follows: 


HAULAGE  COSTS  335 


COST  OF  ELECTRIC  INSTALLATION 

Two  15-ton  locomotives  @  $2300 $4,600 . 00 

One  175-kw.  generator  and  switch-board 2,400.00 

One  McEwen  engine,  18 X 18  in.,  200  hp 2,000 . 00 

Foundations  and  placing  engine  and  generator ....  300 . 00 

Two  72  in.  X 18  ft.  tub.  boilers,  150  hp .,  complete . .  2,800 . 00 

9000  ft.  trolley  wire 1,019 .90 

200  ft.  400,000-cm.  lead  cable  @  55c 110.00 

665  trolley  hangers  @  65c 432.25 

768  bonds  @  35c 268.80 

75  crossbonds  @  35c 26 . 25 

18  interchangeable  trolley  frogs  @  $2 . 75 49 . 50 

1  extra  250-volt  armature 375 . 00 

2  motor  jacks  @  $12.80 25.60 

Extra  fittings  for  motors 86 . 24 

116£  tons  40-lb.  rail  @  $28 . 25,  $3,291 . 13.     Cr.  for 

25-lb.  rails,  $2,056.75 1,234.38 

6055  white-oak  ties  @  lOc 605.50 

65  kegs  4£XHn.  spikes  @  $3.75 244.50 

22  split  switches,  material  and  labor  @  $17.00. .  374.00 

Fish  plates  and  bolts 280.00 

Lumber  for  trolley  supports 76 . 11 

Sundries 3,810.21 

Entire  labor  cost 3,810.21 


Total  of  complete  installation $21,172.79 


COST  OF  MULE  HAULAGE 

• 

Mules,  average  cost $225 . 00 

Mules,  depreciation 20  per  cent 

Mules,  interest. .  < 6  per  cent 


COST  PER  MULE,  DIVIDED  FOR  275  WORK  DAYS,  PER  WORK  DAY 

Depreciation $0. 163 

Interest 0 . 049 

Feed 0.20 

Shoeing  and  stableman .'. 0 . 158 


Total  per  day,  275-day  basis $0 . 57 


336  COAL  MINING  COSTS 


TEAM  HAULAGE,  No.  3  MINE  DAILY  AVERAGE  TONNAGE,  1400  TONS 

17  mules  ©    $0.57 :     $9.69 

9  drivers  ®    2.56 24.24 


Total $33.93 

Team  drivers,  15c.  extra,  or  $1.20  extra  for  8. 
Cost  per  ton  outside  mule  haulage,  2.4c. 

COST  OF  OPERATING  ELECTRICALLY,  275  WORK  DAYS 

2  locomotive  runners  @  $3 . 20 $6 . 40  per  day 

2  trip  riders  @  $2.56 5. 12  per  day 

£  electrician  @  $75  per  month 1 . 08  per  day 

£  fireman  @  $2 . 02 0. 67  per  day 

Fuel,  5  tons  @  75c 3 . 75  per  day 


Total  fixed  labor,  etc $17 . 02  per  day 

Interest  on  investment,  $21,172 . 79  @  6  per  cent $4 . 62 

Depreciation  and  repairs  @  8  per  cent 6.16 

Oil  and  waste : 0 . 30 

Taxes..  0.50 


Total  others $11 . 58 

Total  daily  operating  cost  electrically $28 . 60 

Cost  per  ton  based  on  2000  tons 1.4c 

It  will  be  seen  from  the  foregoing  that  the  plant,  besides 
increasing  the  output  and  saving  Ic.  per  ton,  which  means 
$20  per  day,  practically  pays  for  itself  in  four  years. 

The  following  figures  show  a  comparison  of  cost  previous 
to  1910,  between  mules  and  an  electric  locomotive  at  a  mine 
where  14  mules  were  replaced  by  one  locomotive.  The  output 
of  the  mine  averages  1500  tons  per  day  for  245  working  days 
per  year.  The  cars  weigh  2400  Ib.  empty  and  hold  3600  lb., 
making  a  total  weight  of  6000  lb. 

MULE  HAULAGE 

14  mules  ©  $180  each $2,520.00 

14  sets  harness  @  $25  each 350.00 


$2,870.00 


HAULAGE  COSTS 


337 


INTEREST  AND  DEPRECIATION 


20  per  cent  depreciation  on  $2870. 
6  per  cent  interest  on  $2870 


$574.00 
172.20 


$746.20 


WORKING  EXPENSES  FOR  245  DAYS 

14  mules  feeding,  shoeing,  repairing  harness  and 

care  @  50c.  per  mule,  per  day $1,715.00 

6  drivers  @  $2.80  per  day 4,116.00 

$5,831.00 

$5,831.00 
746.20 

$6,577.20 

Fifteen  hundred  tons  per  day  for  245  days  equals 
367,500  tons  per  year  at  a  haulage  cost  of  1.8c. 
per  ton. 

ELECTRIC  HAULAGE 

Engine,  locomotive,  boiler  and  generator $9,000 . 00 

Switches,  insulators  and  wire 1,200.00 

Cost  of  erecting,  etc 1,000 . 00 

$11,200.00 

INTEREST,  DEPRECIATION,  ETC. 

Interest  at  6  per  cent  on  $11,200 $672.00 

Depreciation  on  boiler,  engine,  etc.  @  9  per  cent.  .  810.00 

Repairs  on  boiler,  engine,  etc.,  @  9  per  cent 810 . 00 

Depreciation  on  switches,  wire,  etc.,  @  5  per  cent.  110.00 

Repairs  on  switches,  wire,  etc.,  @  5  per  cent 110. 00 

$2,512.00 

Engineer  power  house  @  $75  per  month $900 . 00 

Motorman  @  $2.80  per  day 686.00 

Oil  and  waste 100 . 00 

Nipper  on  motor  @  $1 . 50  per  day 367 . 50 

Sand 50.00 

$2,103.50 
From  above $2,512 . 00 

Total $4,615.50 


338  COAL  MINING  COSTS 

Fifteen  hundred  tons  per  day  for  245  days  equals  365,700 
tons  per  year.  This  makes  the  haulage  on  each  ton  of  coal, 
where  electric  locomotives  are  used,  cost  1.27c.  per  ton.  These 
estimates,  taken  from  an  actual  case,  show  a  considerable  dif- 
ference in  favor  of  electric  haulage.  The  cost  of  installing 
mechanical  haulage  is  greater  than  when  a  mine  is  supplied 
with  mules;  however,  when  we  consider  the  cost  of  erecting 
a  stable  and  the  great  loss  due  to  mules  killed  in  accidents, 
the  initial  expenditure  is  not  so  favorable  to  the  use  of  mules. 

Cost  and  care  of  mules. — Frank  Amos,  of  the  Fairmont  Coal 
Co.,  in  1911,  made  the  statement  "that  the  average  life  of 
a  mine  mule  was  3y2  yr.,  and  unless  conditions  were  changed 
to  prolong  life,  the  use  at  the  present  cost  of  the  animal  was 
unprofitable. ' ' 

When  an  animal  of  this  kind  lives  indefinitely  on  the  farm 
it  seems  incredible  that  his  life  should  be  shortened  to  3y2  yr. 
in  the  mine.  There  are  mules  to-day  in  anthracite  mines  that 
have  been  working  20  yr.,  yet  it  is  probable  that  the  aver- 
age life  of  such  animals  in  all  anthracite  mines  does  not  exceed 
5  yr. 

In  purchasing  stock  for  underground  haulage,  activity,  eye- 
sight, feet,  temperament,  strength,  and  wind  are  considera- 
tions, but  if  the  animal  lacks  intelligence  he  has  no  place  in 
the  mine.  In  most  instances  it  is  better  to  deal  with  those  who 
make  a  business  of  furnishing  mules  to  mining  companies,  and 
if  it  is  possible,  to  go  to  the  stock  yards  and  pick  out  the 
animals  rather  than  trust  to  the  dealer's  judgment.  This  sug- 
gestion is  made  because  the  dealer  scarcely  knows  more  about 
the  animals  than  the  purchaser,  and  does  not  know  the  con- 
ditions under  which  the  animals  must  work. 

Many  animals  which  act  and  look  right  on  the  surface  are 
most  unsatisfactory  underground,  for  which  reason  the  sug- 
gestion is  made  that  after  the  animals  are  picked  out  an  agree- 
ment should  be  made  with  the  dealer  that  in  case  any  of  them 
do  not  act  rightly  in  the  mine  they  may  be  exchanged. 

While  a  mule's  heels  are  not  to  be  trusted,  nevertheless 
the  humane  driver  and  his  mule  become  good  chums.  If,  after 
a  3-day  training  period  the  mule  does  not  appear  active  on 
his  feet  and  to  use  judgment,  he  should  be  taken  from  the 


HAULAGE  COSTS  339 

mine,  as  he  is  unsuited  to  the  work  and  cannot  be  depended 
upon  to  look  out  for  himself  when  occasion  demands. 

According  to  two  authorities  the  animals  should  be  fed  as 
follows:  First,  hay,  next  water,  and  then  grain.  Animals 
should  eat  hay  at  least  half  an  hour  before  being  given  grain. 
If  the  water  is  given  last  it  washes  the  food  into  the  intestines 
before  it  is  acted  upon  by  the  gastric  juices.  If  the  hay  is  given 
after  the  grain  it  carries  the  grain  with  it,  for  the  hay  is  prin- 
cipally digested  in  the  intestines,  while  the  grain  is  acted  upon 
by  the  stomach  for  the  most  part. 

Corn  is  richer  in  fat  than  oats ;  therefore,  for  strength,  feed 
corn,  and  for  speed,  feed  oats.  For  an  illustration,  race  horses 
are  fed  oats,  and  the  experienced  teamster  will  favor  the  feed- 
ing of  corn. 

Dr.  I.  C.  Newhard,  Chief  Veterinarian  of  the  Philadelphia 
&  Beading  Coal  and  Iron  Co.,  in  1911,  experimented  with 
various  feeds  and  found  that  two-thirds  crushed  oats  and  one- 
third  cracked  corn  the  most  reliable.  "A  handful  of  coarse 
ground  pure  salt  should  be  fed  to  each  mule  twice  a  week.': 
Dr.  Frank  Amos,  who  is  in  West  Virginia,  suggests  a  coarse- 
crushed  feed,  about  two-thirds  corn  and  one-third  oats. 

Mine  stock  will  consume  about  12  Ib.  per  head  per  day 
of  this  feed  and  about  15  Ib.  of  hay.  If  a  horse  or  a  mule 
has  not  cleaned  up  its  former  feed  the  troughs  should  be 
cleaned  and  less  put  in  the  next  time,  until  it  is  ascertained 
just  how  much  it  takes  to  keep  them.  The  animal  should  have 
about  all  it  will  eat,  but  it  is  better  to  give  not  quite  enough 
than  too  much.  Too  much  grain  will  cause  acute  indigestion, 
paralyze  the  walls  of  the  stomach,  and  usually  results  in  death. 
A  stable  boss  will  make  a  great  mistake  by  feeding  too  much 
and  allowing  food  to  stand  before  the  animals  all  the  time. 
While  this  method  will  increase  flesh  for  a  short  period  the 
animals  eventually  break  down  through  their  digestive  organs 
being  destroyed. 

Grain  should  not  be  placed  in  the  animals*  troughs  ready 
for  them  when  they  come  in  from  work,  and  it  is  better  for 
them  to  be  without  grain  at  noontime  than  to  be  without  water, 
but  by  all  means  give  them  three  feeds  a  day.  Plenty  of  water 
will  keep  the  digestive  organs  in  good  condition,  while  large 
quantities  of  grain  and  no  water  will  destroy  them. 


340  COAL  MINING  COSTS 

To  give  a  feed  of  good  bran  once  a  week  will  aid  the  con- 
ditioning of  stock,  keep  the  bowels  open,  and  reduce  fever, 
which  is  caused  by  strong  grain. 

The  Fairmont  Coal  Co. 's  records  for  1905  show  that  26  per 
cent  of  their  stock  either  died,  was  killed,  or  had  to  be  dis- 
posed of  at  practically  nothing,  on  account  of  being  crippled 
and  worn  out.  There  is  probably  no  part  of  the  company's 
business  in  which  the  loss  is  so  great,  and  one-half  of  this  is 
brought  about  by  carelessness  and  neglect.  There  is  probably 
no  other  business  that  requires  the  use  of  stock  in  which  the 
loss  is  so  great,  and  this  in  face  of  the  fact  that  the  facilities 
furnished,  with  the  exception  possibly  of  good  roads  inside 
the  mines,  are  the  best  that  can  be  had. 

The  maintaining  of  live  stock  is  no  little  item,  and  in  cases 
there  is  an  average  of  5  per  cent  of  the  total  number  of  animals 
standing  in  the  barns  all  the  time  unfit  for  service  on  account 
of  having  been  crippled.  The  feed  for  this  stock,  besides  other 
expenses  and  the  loss  of  their  work,  cost  one  company  in  1911, 
over  $6000  per  year. 

Where  mine  stock  is  given  good  attention,  the  upkeep  is 
reduced  to  a  minimum,  more  work  is  obtained,  and  the  animals 
are  more  valuable. 

Professor  Ihlseng  estimates  the  capacity  for  work  of  the 
mine  mule  on  a  level  track  as  between  30  and  80  gross  ton- 
miles  in  8  hr.  and  the  work  of  the  average  mule  for  the  same 
time  as  between  40  and  50  gross  ton-miles  (5  to  6  ton-miles  per 
hour)  with  a  limiting  grade  of  3  per  cent.  Hughes  gives 
examples  varying  all  the  way  from  3  to  16  ton-miles  per  hour 
on  level  track.  Other  conditions  being  equal  the  condition 
that  really  determines  the  work  that  a  mine  mule  can  do  is 
the  length  of  the  haul,  since  a  large  part  of  the  time  is  con- 
sumed in  changing  and  waiting  for  trips.  The  average  of  a 
number  of  actual  working  conditions  shows  that  4  to  6  ton- 
miles  per  hour  is  a  good  average  for  the  work  of  an  ordinary 
mine  mule,  in  a  flat  seam,  under  usual  mine  conditions,  though 
this  will  increase  slightly  with  the  length  of  haul.  With  a 
normal  load  the  limiting  grade  under  which  a  mule  can  bo 
worked  to  an  advantage  is  about  3  or  4  per  cent.  As  loaded 
mine  cars  will  slide  on  iron  rails,  with  4  sprags,  on  grades 


HAULAGE  COSTS 


341 


varying  from  6  to  8  per  cent  the  safe  down  grade  for  a  mule 
to  work  on  is  limited  to  this. 

Figures  gathered  from  a  number  of  years  experience  in  a 
level  seam  5l/2  ft.  thick  show  that  a  good  2-mule  team,  with 
an  efficient  driver  and  helper,  will  handle  from  50  to  60  cars 
of  2y2  tons  capacity  in  a  10-hr,  shift  when  the  haul  does  not 
exceed  2000  ft.  Allowing  two  hours  lost  time  and  assuming 
the  weight  of  the  empty  car  to  be  1^  tons  each  mule  has 
accomplished  7  ton-miles  of  work  per  hour. 

An  additional  expense  that  must  be  allowed  for  in  animal 
haulage  in  low  seams  is  the  cost  of  brushing  to  obtain  suffi- 
cient headroom.  Haulage  expenses  have  been  increased  as 
much  as  3c.  per  ton  (about  1912)  from  this  cause  and  the 
decreased  efficiency  of  the  animals. 


I    I    I    I    I    I    I    I    I  I    I  I  I 

-A  LOCOMOTIVE,!  EN6INEER.2SRAKEMEN  $2442  PER  DAY - 

///,        "     !      7      "          1BPAKEMAN  20.07  »  » 

\-\-JDRIVER,  2 RUNNERS,  3 MULES  17.60  »  » — 

\\1     »     I     1  RUNNER,  3    »  13.24  »  » 

'  v  *            !     /       »           Z     »  11.74  »  »_ 

\..-l  DRIVER,    S     »  7.37  n  » 

JMULE  5.67  »  ». 


Q  0.10  0.10  0.30  0.40  050  060  0.70  0.60  090  1.00  1.10  1ZO  1.30  1.40  150  1.60  1.70  IflO 
Cost    of     Car  (^c.v    C6Jr). 

FIG.  22. — Chart  showing  relative  economy  of  mule  and  motor  haulage. 

Additional  ventilation  must  also  be  provided  where  animal 
haulage  is  used,  most  state  mining  laws  providing  that  500 
cu.  ft.  per  min.  be  allowed  for  each  animal  in  the  mine.  In 
a  mine  using  50  mules  this  would  mean  that  an  additional 
25,000  cu.  ft.  of  air  per  minute  be  provided. 

The  working  life  of  the  average  mule  is  seven  to  ten  years. 
Because  of  the  hard  working  conditions  and  generally  fre- 
quent accidents  it  is  necessary  to  keep  a  good  reserve  supply 
of  animals  available  which  increases  the  investment,  as  well  as 
costs  of  maintenance. 

The  chart  shown  in  Fig.  22,  will  give  a  quick  approximate 
solution  of  the  relative  economy  of  mule  and  motor  haulage 
where  the  conditions  are  known.  Other  factors,  such  as  availa- 
bility of  power  and  the  possibility  of  securing  locomotives  and 


342 


COAL  MINING  COSTS 


equipment  are  matters  for  consideration  after  or  before  the 
relative  costs  of  the  different  organizations  are  determined. 
The  curves  are  made  up  from  figures  similar  to  those  appear- 
ing in  Table  I: 

TABLE  I 
COST  PER  CAR  FOR  ONE  DRIVER  AND  ONE  MULE 


Cars 

Cost 

Cars 

Cost 

Cars 

Cost 

Cars 

Cost 

4 

$1.47 

8 

$0.73 

16 

$0.37 

30 

$0.195 

5 

1.17 

12 

.49 

18 

.33 

40 

.147 

6 

.98 

14 

.42 

21 

.28 

60 

.098 

These  when  plotted  give  the  basis  for  the  regular  curve  for  an 
organization  or  transportation  unit  of  one  driver  and  one  mule. 
The  other  curves  shown  are  plotted  in  a  like  manner.  In  the 
case  of  locomotive  haulage,  the  power,  installation  and  repair 
costs  given  or  estimated  for  any  single  locality  will  not  be  cor- 
rect for  every  installation.  This  is  one  weak  point  in  the  tabu- 
lation. However,  because  unforeseen  conditions  usually 
develop  in  mine  electric  installations,  it  is  best  to  be  on  the  safe 
side  and  figure  locomotive  maintenance  high. 

A  close  inspection  of  these  curves  will  also  show  that  tho 
handling  of  four  or  five  more  cars  makes  much  more  difference 
in  the  cost  per  car  than  does  a  few  cents  in  the  figures  repre- 
senting power,  repairs,  etc.  For  instance,  in  the  case  of  a 
crab  locomotive  with  one  engineer  and  two  brakemen  working 
in  pitching  rooms,  let  us  estimate  that  this  machine  will  handle 
40  cars  per  day.  This,  from  Fig.  22,  would  cost  60c.  per  car. 
If  another  place  can  be  added  so  that  the  machine  will  handle 
44  cars,  it  will  do  so  at  a  cost  of  56c.  per  car — a  saving  of  four 
cents  per  car,  or  $1.76  per  day.  We  cannot  always  estimate 
mine  haulage  possibilities  closer  than  four  cars,  but  our  cost 
figures  certainly  will  not  be  in  error  to  the  amount  of  $1.76. 

Two  representative  problems  are  given  below  to  illustrate 
the  use  of  the  curves : 

Problem  1 — Assumption  by  colliery  superintendent  that  his 
transportation  gathering  cost  shall  not  be  over  60c.  per  car. 


HAULAGE  COSTS  343 

How  many  cars  must  each  unit  handle  in  order  to  give  him 
this  cost? 

Follow  the  dash  line  on  Fig.  22:  It  crosses  one  driver  and 
one  mule  at  10  cars;  one  driver  and  two  mules  at  12  cars;  one 
driver,  one  runner  and  two  mules  at  19  cars;  one  driver,  one 
runner  and  three  mules  at  22  cars ;  one  driver,  two  runners  and 
three  mules  at  29  cars;  one  locomotive,  one  engineer  and  one 
brakeman  at  34  cars ;  one  locomotive,  one  engineer  and  two 
brakemen  at  40  cars. 

Locomotive  Data:  Per  Day 

Power,  $90  per  month $3.60 

Installation,  $2000,  5  per  cent  of  $2000  1  $2200 

Depreciation — Locomotive,  15%  of  $8000 
Repairs — Locomotive  lines,  etc.,  $900  per  year  J  7.33 

Labor — Engineer  at  $0.28,  plus  $0.25  allowance  for  9  hrs.   4 . 77 
Brakeman  at  $0.2351,  plus  $0.25  allowance  for 

9  hours 4.36 

Mule  Data: 

Mule  value,  $250;  20%  depreciation;  5%  interest 0.23 

Feed  and  care 1 .25 

Labor— Driver  at  $0.2351,  plus  $0.25  allowance. 
Runner  at  $0.2351,  plus  $0.25  allowance. 

One  driver,  two  runners  and  three  mules  in  a  certain  sec- 
tion where  there  are  pitching  rooms  will  handle  25  cars  per 
day.  This  section  due  to  its  development  can  produce  50  cars 
per  day  with  two-mule  organizations.  Which  relatively  is  the 
cheaper  plan,  to  use  two-mule  organizations  or  one  crab  loco- 
motive to  handle  50  cars? 

By  following  the  two  dotted  lines  in  Fig.  22  it  will  be  seen 
that  one  driver,  two  runners  and  three  mules  will  handle  25 
cars  at  a  cost  of  70c.  per  car ;  one  locomotive,  one  engineer  and 
two  brakemen  will  handle  50  cars  at  a  cost  of  48c.  per  car.  In 
this  case,  of  course,  the  locomotive  is  the  better  installation. 

Close  inspection  of  the  curves  in  Fig.  22  will  also  show  that 
there  is  a  certain  fairly  definite  limit  to  the  number  of  cars 
handled,  below  which  the  cost  per  car  increases  rapidly.  This 
limit  can  be  taken  from  the  curves  by  noting  the  point  at  which 
the  curve  from  right  to  left  deviates  from  practically  a  straight 
line.  Thus,  for  instance,  in  the  case  of  the  one  driver  and  one 
mule  curve,  the  line  is  practically  straight  from  the  right  hand 


344 


COAL  MINING  COSTS 


side  of  the  figure  to  the  10  or  12  car  point,  and  from  here 
turns  rapidly  into  a  curve.  This  possibly  is  better  illustrated 
by  Table  II. 

Taking  the  limit  for  the  differences  or  decrease  in  cost  per 
car  at  4c.,  we  see  that  12  cars  must  be  handled  by  one  driver 
and  one  mule,  18  cars  by  one  driver,  one  runner  and  three 
mules,  etc.,  before  this  limit  is  reached.  Taking  2c.  as  the 
limit  16  cars  must  be  handled  by  one  driver  and  one  mule,  and 
a  corresponding  number  by  other  units  before  reaching  this 
limiting  figure. 


TABLE  II 

COST  PER  CAR  FOR  DIFFERENT  UNITS  FOR  INCREASING  NUMBER  OF  CARS 


. 

1 

. 

.2 
fc£« 

I 

£§*1 

c  §  2 

1 

'II 

I 

•g! 

1 

Ill 

• 

V 

QtfS 

0> 

Ill 

HI 

I 

III 

| 

Q 

1 

2 

3 
4 
6 

6 
7 
8 
9 
10 
11 
12 
13 
14 
15 
16 
17 
18 
19 
20 
21 
22 
23 
24 
25 
26 
27 
28 

5.87 
2.93 
1.96 
1.47 
1.17 
.98 
.84 
.74 
.65 
.59 
.53 
.49 
.45 
.42 
.39 
.37 
.35 
.33 
.31 
.29 
.28 
.27 
.26 

2.94 
.97 
.49 
.30 
.19 
.14 
.10 
.09 
.06 
.06 
.04 
.04 
.03 
.03 
.02 
.02 
.02 
.02 
.02 
.01 
.01 
.01 

7.37 
3.68 
2.46 
1.84 
1.47 
1.22 
1.05 
.92 
.82 
.74 
.67 
.61 
.57 
.53 
.49 
.46 
.43 
.41 
.39 
.37 
.35 
.33 
.32 

3.69 
1.22 
.62 
.37 
.25 
.17 
.13 
.10 
.08 
.07 
.06 
.04 
.04 
.04 
.03 
.03 
.02 
.02 
.02 
.02 
.02 
.01 

11.74 

5.87 
3.91 
2.94 
2.35 
1.96 
1.68 
1.47 
1.31 
1.17 
1.07 
.98 
.91 
.84 
.78 
.73 
.69 
.65 
.62 
.59 
.56 
.53 
.51 
.49 
.47 
.45 
.43 
.42 

5.87 
1.96 
.97 
.59 
.39 
.28 
.21 
.16 
.14 
.10 
.09 
.07 
.07 
.06 
.05 
.04 
.04 
.03 
.03 
.03 
.03 
.02 
.02 
.02 
.02 
.02 
.01 

13.24 
6.62 
4.41 
3.32 
2.65 
2.21 
1.89 
1.65 
1.47 
1.32 
1.20 
1.10 
1.02 
.95 
.88 
.83 
.78 
.74 
.70 
.66 
.63 
.60 
.57 
.55 
.53 
.51 
.49 
.47 

6.62 
2.21 
1.09 
.67 
.44 
.32 
.29 
.18 
.15 
.12 
.10 
.08 
.07 
.07 
.05 
.05 
.04 
.04 
.04 
.03 
.03 
.03 
.02 
.02 
.02 
.02 
02 

17.60 
8.80 
5.87 
4.40 
3.52 
2.94 
2.52 
2.20 
1.96 
1.76 
.60 
.46 
.35 
.26 
.17 
.10 
.03 
.98 
.93 
.88 
.84 
.80 
.76 
.73 
.70 
.67 
.65 
.63 

8.80 
2.93 
1.47 
.88 
.58 
.42 
.32 
.24 
.20 
.16 
.14 
.11 
.09 
.09 
.07 
.07 
.05 
.05 
.05 
.04 
.04 
.04 
.03 
.03 
.03 
.02 
.02 

20.07 
10.03 
6.69 
5.02 
4.01 
3.35 
2.87 
2.51 
2.23 
2.01 
1.83 
1.67 
1.54 
1.44 
1.34 
1.25 
1.18 
1.11 
1.05 
1.00 
.96 
.92 
.88 
.84 
.80 
.77 
.74 
.72 

10.04 
4.34 
1.67 
1.01 
.66 
.48 
.36 
.28 
.22 
.18 
.16 
.13 
.10 
.10 
.09 
.07 
.07 
.06 
.05 
.04 
.04 
.04 
.04 
.04 
.03 
.03 
.03 

24.42 
12.21 
8.14 
6.11 
4.89 
4.07 
3.49 
3.06 
2.72 
2.44 
2.22 
1.04 
1.88 
1.74 
1.63 
1.52 
1.44 
1.36 
1.28 
1.22 
1.16 
1.11 
1.06 
1.02 
.98 
.94 
.90 
.87 

12.21 
4.07 
2.03 
1.22 
.82 
.58 
.43 
.34 
.28 
.22 
.18 
.16 
.14 
.11 
.11 
.08 
.08 
.08 
.06 
.06 
.05 
.05 
.04 
.04 
.04 
.04 
.03 

HAULAGE  COSTS  345 

The  following  are  the  results  obtained  through  the  instal- 
lation of  five  storage-battery  locomotives  in  the  Red  Ash  vein 
at  Exeter  Colliery  of  the  Lehigh  Valley  Coal  Co.* 

In  the  fifth  vein,  one  locomotive,  operating  from  inside  slope 
to  chambers,  gangways  and  airways  on  roads  Nos.  1002  and 
1006,  with  chambers  pitching  both  ways  and  grades  in  some 
places  as  high  as  9  per  cent,  handles  fifty  cars  with  a  maxi- 
mum run  of  about  1800  ft.  When  replacing  the  empty  cars 
this  locomotive  is  assisted  by  a  second  locomotive.  The  latter 
also  covers  chambers,  airways  and  gangways  on  roads  Nos. 
1001  and  1006  and  handles  sixteen  cars  with  a  run  of  1300 
ft.  in  each  gangway  with  grades  ranging  as  high  as  7  per 
cent.  To  replace  these  two  motors  with  mule  power  would 
take  fifteen  mules,  five  drivers  and  five  runners. 

In  the  Babylon  vein,  one  motor  hauls  from  the  various 
working  faces  on  roads  Nos.  144  and  147  and  handles  twenty- 
five  cars  daily,  having  a  maximum  run  of  2100  ft.,  and  delivers 
coal  to  the  head  of  No.  9  plane.  It  would  take  nine  mules, 
three  drivers  and  three  runners  to  replace  this  motor. 

The  fourth  locomotive  collects  thirty-two  cars  from  roads 
Nos.  39,  46  and  50  and  delivers  coal  to  a  big  turnout,  having 
a  maximum  run  in  each  road  of  3500  ft.,  2700  ft.,  and  3000  ft. 
respectively.  Nine  mules,  three  drivers  and  three  runners 
would  be  required  to  replace  this  motor. 

The  fifth  locomotive,  operating  as  a  collecting  locomotive 
on  road  No.  5  and  in  working  faces  between  chambers  Nos.  33 
and  47,  and  assisting  in  concentrating  coal  from  roads  Nos. 
65  and  4,  handles  thirty  cars  daily  over  a  run  of  1000  ft. 
having  grades  up  to  5  per  cent  against  loaded  trips.  It  would 
require  six  mules,  two  drivers  and  two  runners  to  replace  this 
motor. 

As  compared  with  mule  haulage  there  is  ease  on  record 
in  the  Ohio  fields  where  one  storage  battery  6-ton  motor  re- 
placed 12  mules  and  four  drivers,  handling  85  to  100  cars  on 
grades  up  to  as  high  as  17  per  cent  it  being  necessary  to  sand 
the  track  on  the  steeper  grades  of  course.  The  economy 
effected  in  this  case,  after  generous  allowances  for  deprecia- 
tion, is  obvious. 

*  Extract  from  "The  Storage  Battery  Locomotive  for  Gathering  Pur- 
poses" in  Employees  Magazine  of  the  Lehigh  Valley  Coal  Co. 


346  COAL  MINING  COSTS 

Gasoline  motor  vs.  animal  haulage. — The  Long  Branch  Coal 
Co.  installed  a  gasoline-motor  haulage  system  at  its  mine  in  West 
Virginia,  that  materially  reduced  haulage  costs  and  at  the 
same  time  increased  the  output.  The  supplanting  of  the  old 
system  of  mule  haulage  was  done  in  1913  with  the  idea  of 
cutting  down  operating  costs.  The  company  purchased  a 
7-ton  gasoline  locomotive  which  was  put  into  service  in  the 
summer  of  1913.  At  the  end  of  the  month  of  August,  a  com- 
parison was  made  with  the  month  of  February,  the  most  pro- 
ductive month  during  the  regime  of  mule  haulage.  A  summary 
of  this  comparison  is  given  below : 

COMPARATIVE  COSTS  OF  THE  Two  SYSTEMS  OF  HAULAGE 

Total  cost  of  mule  haulage,  per  month $810 . 00 

Total  cost  of  gasoline  haulage,  per  month . . . ." 529 . 63 


Decrease  in  haulage  cost,  per  month $280 . 37 

Total  coal  tonnage  by  gasoline  locomotive,  tons 11,601 

Total  coal  tonnage  by  mules,  tons 7,848 


Increase  in  coal  output,  tons 3,753 

The  analysis  of  the  above  summary  is  shown  by  the  follow- 
ing detail  comparison  between  the  cost  of  haulage  by  mules 
and  by  a  combination  of  gasoline  locomotive  and  mules  for 
gathering. 

Mule  Haulage — Month  of  February,  1913 

Length  of  haul  one  way,  feet 2,000 

Maximum  grade  against  loads,  per  cent 5 . 625 

Average  grade  against  loads,  per  cent 3.0 

Total  tonnage  per  month  of  24  days 7848 

15  mules — feed  and  upkeep  per  day  @  60c $  9 . 00 

11  drivers — wages  per  day  @  $2 . 25 24 . 75 


$33.75 

24  working  days  @  $33 . 75— cost  per  month $810 . 00 

Total  haulage  cost  per  ton  of  coal 0 . 103 

Total  haulage  cost  per  ton-mile 0 . 272 

Gasoline  Locomotive  Haulage  in  Connection  with  Gathering  by  Mules — Month 

of  August,  1913 

Length  of  haul  one  way,  feet 3000 

Maximum  grade  against  loads,  per  cent 5 . 625 

Average  grade  against  loads,  per  cent 3.0 

Total  tonnage  per  month  of  25  days 11,601 

Expense  of  mules  and  drivers  for  gathering,  25  working 

days $313.20 


HAULAGE  COSTS  347 


COST  OF  OPERATING  LOCOMOTIVE,  25  DAYS 

1  motorman  @  $3  per  day $75 . 00 

1  trip  rider  @  $2  per  day 50 . 00 

300  gal.  gasoline  @  20c.  per  gal 60 . 00 

38  gal.  engine  oil  @  40c.  per  gal 15 .20 

3  gal.  black  oil  @  16c.  per  gal 0.48 

Cup  grease 0 . 50 

Waste 0.25 

Repairs  on  motor 15 . 00 


Total  operating  expense  of  locomotive $216 . 43 


Total  haulage  cost  per  month $529 . 63 

Total  haulage  cost  per  ton  of  coal 0 . 0456 

Total  haulage  cost  per  ton-mile 0 . 08.03 

Cost  per  ton  of  coal — mule  haulage 0 . 103 

Cost  per  ton  of  coal — locomotive  haulage 0.0456 


Saving  per  ton  of  coal $0 . 0574 

Cost  per  ton-mile — mule  haulage 0 . 272 

Cost  per  ton-mile — locomotive  haulage 0 . 0803 


Saving  per  ton  mile $0 . 1917 

On  a  basis  of  12  months  the  cost  by  mule  haulage  for 

one  year  ($810X12) $9720.00 

By  locomotive  for  one  year  ($529  X 12) 6348 . 00 


Yearly  saving $3372.00 

With  the  gasoline  haulage  system  the  tonnage  of  the  mine 
has  been  increased  on  an  average  of  25  per  cent  per  month, 
and  the  company  has  dispensed  with  6  double  teams  or  12 
mules  and  6  drivers,  a  total  monthly  saving  in  expense  of 
$496.80. 

The  management  of  the  Midvalley  Coal  Co.  in  Pennsyl- 
vania, in  1911,  effected  a  saving  of  32.2  per  cent  on  coal  hauled 
by  substituting  the  gasoline  locomotive  for  mules. 

This  comparison  seems  too  conservative,  because  the  loco- 
motive was  in  a  position  where  its  full  capacity  could  not  be 
demonstrated,  in  fact  was  idle  a  large  part  of  the  time,  mak- 
ing only  24  miles  per  day  when  it  is  capable  of  doing  more 
than  twice  as  much.  The  management  estimates  that  a  second 
locomotive  that  was  to  be  installed  would  save  practically  50 
per  cent  over'  the  present  system  of  haulage  as  it  was  to  be 


348  COAL  MINING  COSTS 

placed  on  a  level  and  have  sufficient  work  to  keep  it  moving, 
displacing  15  mules. 

By  comparing  the  cost  of  haulage  with  the  gasoline  loco- 
motive at  Midvalley  and  the  average  cost  of  electric  locomotive 
haulage  as  furnished  in  the  Coal  and  Metal  Miners'  Pocket- 
book,  it  was  found  that  there  was  a  lessened  cost  of  27.9  per 
cent  in  favor  of  gasoline.  The  locomotive  uses  naphtha  for 
fuel,  it  being  less  dangerous  and  better  to  handle  than  gasoline. 
The  consumption  of  naphtha  is  about  15  gal.  per  day  ($1.50 
per  day),  where  if  gasoline  were  used  the  consumption  would 
be  about  12  gal.  for  the  same  work  while  the  cost  would  be 
30c.  more. 

The  Midvalley  locomotive,  rated  as  a  9-ton  locomotive,  has 
.about  the  following  dimensions :  Length,  150  in. ;  width,  59 
in. ;  height,  60  in. ;  wheel  base,  48  in. ;  and  diameter  of  driving- 
wheels,  24  in.  The  following  table  gives  the  detail  of  the 
average  work  performed  daily  by  the  first  locomotive  in  6 
months,  during  which  period  approximately  2  hr.  out  of  a 
9-hr,  day  were  devoted  to  switching,  a  feature  which  fails  to 
show  on  the  cost  sheet: 

Average  tonnage  of  loaded  cars  per  day 550  tons 

Average  tonnage  of  empty  cars  per  day 250  tons 

Average  mileage  of  loaded  cars  per  day 12  miles 

Average  mileage  of  empty  cars  per  day 12  miles 

Weight  of  one  loaded  car 5£  tons 

Weight  of  one  empty  car 1\  tons 

Average  number  of  cars  per  train 8  cars 

COST  OF  OPERATION 

Wages  of  operator  and  helper  per  day $3 . 35 

Cost  of  fuel  per  day 1 . 50 

Cost  of  lubricating  oils  per  day .12 

Consumption  of  fuel  in  gallons  daily 15 

MAINTENANCE 

Cost  of  maintenance  of  locomotive  for  six   months, 

including  repairs  and  labor $65 . 14 

The  substitution  of  gasoline-motor  haulage  for  mules  at 
some  mines  near  Rockwood,  Tenn.,  in  1911,  presented  some 
interesting  figures  on  the  comparative  costs  of  these  two 


HAULAGE  COSTS  349 

methods  of  haulage.  The  coal  in  this  mine  is  collected  on  side- 
tracks on  the  main  entries  by  mules  or  by  rope  from  cross- 
entries,  the  nearest  parting  being  iy2  miles  from  the  slope. 
Mule  haulage  had  been  used  on  this  long  haul  and  this  had 
been  found  so  expensive  that  it  was  decided  to  install  three 
gasoline  motors.  The  total  output  of  the  mine  was  between 
600  and  700  tons  a  day  all  of  which  passed  over  this  long  main- 
haul  distances  varying  from  1%  to  2  miles. 

All  the  extra  work  in  the  mines,  necessary  for  the  instal- 
lation of  these  motors,  was  some  slight  trimming  of  the  rib 
and  top  in  places,  so  as  to  give  ample  clearance  for  the  motors 
and  going  over  the  track  to  replace  with  20-lb  rail,  the  places 
on  the  entry  where  a  lighter  rail  had  heretofore  been  used. 

There  was  no  difficulty  found  by  reason  of  the  many  curves, 
as  the  motors  have  a  4-ft.  wheel  base,  and  can  take  a  curve 
of  25-ft.  radius.  The  locomotives  are  6  tons  each,  and  were 
built  for  the  mine  gage  of  33  in.  They  are  designed  with 
4-cylinder  engines,  of  ample  power  to  slip  the  wheels,  and  all 
parts  are  well  protected,  as  is  necessary  for  mine  use. 

The  mine  cars  used  are  about  1400  Ib.  in  weight,  and  carry 
!1/5  tons  of  coal.  As  the  grade  is  in  favor  of  the  loads,  the 
empty  cars  up  the  entry  make  the  load  for  the  motor.  The 
regular  20-car  trips  are  handled  without  difficulty,  and  on  trial 
trips  40  cars  have  been  taken  up  the  entry. 

These  three  motors  replaced  23  mules.  The  comparative 
estimate  of  mule  and  motor  haulage  on  one  entry  was  as 
follows : 

10  twenty  car  trips  equals 224  tons 

By  mules: 

4  drivers,  at  $1.65 $6.60 

9  mules,  at  50c 4.50        $11.10 


By  motor: 

1  motorman,  per  day $2 . 05 

1  coupler,  per  day 1 . 65 

13  gal.  gasoline,  at  ll^c 1 . 50 

2  Ib.  carbide,  at  4c 08 

\  gal.  gasoline  engine  oil,  at  23c 12 

1  gal.  transmission  case  oil 24        $  5.64 


Saving  by  motor $  5 . 46 

Or.  49 . 1  per  cent 


350  COAL  MINING  COSTS 

These  motors  use  12  to  13  gal.  of  gasoline  each,  per  shift. 
The  Connellsville  Central  Coke  Co.  converted  from  horse  to 
gasoline  motor  haulage  in  1915  and  comparison  of  the  results 
obtained  are  of  value.  Twenty-nine  horses  had  formerly  hauled 
the  output  of  700  cars,  to  the  haulage  rope  or  to  the  shaft 
bottom  as  the  case  might  be.  The  coal  from  some  of  the  flats 
was  handled  independently  of  the  haulage  rope.  This  repre- 
sented a  net  tonnage  of  only  41.5  tons  per  horse,  the  tonnage 
per  unit  being  low,  not  only  because  of  the  excessive  length 
of  haul,  but  because  of  the  heavy  grades,  which  averaged 
6.5  per  cent  in  the  butt  headings.  The  situation  evidently 
needed  corrective  treatment. 

Four  hundred  mine  cars  are  used  in  hauling  the  coal.  These 
are  44-in.  track  gage  and  have  a  capacity  of  approximately 
4000  Ib.  per  car,  which  weigh  when  empty  about  2000  Ib.  As 
the  daily  output  is  1700  tons,  it  is  about  twice  the  capacity  of 
the  mine  cars. 

On  all  main  haulageways  40-lb.  rail  is  used,  and  on  the 
flats  and  subsidiary  butts  the  rails  weigh  25  Ib.  per  yard.  The 
joints  are  all  fishplated,  and  the  track  is  well  ballasted  and 
carefully  graded. 

A  5-ton  gasoline  locomotive  was  selected  and  put  in  opera- 
tion in  September,  1914.  It  pulls  15-car  trips  in  No.  5  flat  right 
and  20-car  trips  in  F  flat  on  the  left,  the  latter  being  a  one- 
way haulage  of  approximately  2000  ft.  and  the  other  being 
roughly  half  that  length.  The  cars  are  delivered  to  the  rope 
at  the  main  butt  entry. 

The  grades  on  both  headings  are  partly  in  favor  and  partly 
against  the  load.  Thus  in  the  No.  5  flat  there  is  a  grade  which 
averages  1%  per  cent  against  the  load  extending  for  the  whole 
distance  between  two  of  the  butt  entries,  and  in  F  flat  there 
is  a  grade  averaging  about  1.2  per  cent  against  the  load  and 
nearly  1200  ft.  long.  It  is  easy  to  see  that  conditions  more 
favorable  might  have  been  chosen.  The  operating  expenses  are 
as  follows: 

Locomotive  runner  per  9-hour  day $2 . 75 

Trip  rider  per  9-hour  day 2.60 

11  gal.  of  gasoline  at  12c.  per  gal 1 . 32 

J  gal.  lubricating  oil  at  22c.  per  gal 06 

Cup  grease,  waste,  oil,  etc .05 


Total.  .  $6.78 


HAULAGE  COSTS  351 

The  results  obtained  from  this  locomotive  were  so  satisfac- 
tory that  a  duplicate  was  purchased  and  shipped  to  the  mine 
in  January,  1915.  This  locomotive  is  working  in  C  flat  on  a 
haul  which  measures  2700  ft.  one  way  and  in  D  flat  where  the 
haul  is  3000  ft.  one  way.  Both  flats  have  grades  favorable  to 
the  load.  However,  in  D  flat  there  is  a  grade  about  300  ft. 
long  which  runs  about  1.1  per  cent  against  the  load.  At  present 
each  of  these  locomotives  is  hauling  about  250  cars  per  day, 
and  handling  about  500  tons.  It  is  expected  that  when  certain 
grades  are  made  more  even  and  when  delays  are  eliminated 
300  cars  loaded  with  600  tons  will  be  handled  by  each  unit. 

For  the  months  of  April  and  May,  1915,  the  locomotive  in 
F  flat  and  No.  5  right  averaged  260  cars  daily;  on  this  basis 
the  comparative  haulage  costs  per  day  are  as  follows : 

Horse  haulage: 

11  drivers  at  $2.60 $28.60 

Feeding  11  horses  at  50c.  each 5 . 50 


$34.10 
Motor  haulage: 

6  drivers  at  $2.60 $15.60 

Feeding  6  horses  at  50c.  each 3 . 00 

1  motorman 2 . 75 

1  snapper 2 . 60 

Gasoline,  oil,  grease,  etc 1 . 51 

25.46 

Savings  per  day  accomplished  by  use  of  gasoline 
motor $  8.64 

The  locomotives  when  fully  equipped  weigh  10,500  Ib. 
Over  all  they  are  144  in.  long,  55  in.  wide  and  46  in.  high  and 
their  wheels  are  18  in.  in  diameter.  They  are  each  equipped 
with  5  X  6-in.  four-cylinder  four-cycle  engines  of  the  vertical 
type,  specially  designed  for  mine-locomotive  service  and  cap- 
able of  delivering  25  hp.  to  the  wheels  at  800  r.p.m. 

The  volume  of  tonnage  is  the  principal  determining  factor  in 
deciding  when  to  substitute  motor  for  animal  haulage.  Care 
must  be  exercised  not  to  make  a  change  before  the  motor  be- 
comes the  most  economical  method.  It  is  not  possible  to  prescribe 
exact  limitations  for  animal  haulage  but  as  a  general  rule  where 
the  haul  exceeds  one-half  mile  or  where  the  cost  of  hauling  300 


352 


COAL  MINING  COSTS 


tons  on  the  main  haulage  way  amounts  to  $1800  a  year  (these 
figures  as  of  1910)  it  is  usually  economy  to  install  some  form 
of  motor  haulage. 

A  comparison  between  gasoline  motor  and  mule  haulage 
was  compiled  at  the  mines  of  the  Shade  Coal  Mining  Co.  in 
Pennsylvania,  the  results  of  which  are  given  herewith.  The 
coal  is  handled  in  185  mine  cars  of  2500  Ib.  capacity  and  weigh- 
ing 1000  Ib.  each.  The  outside  and  main  entry  haulage  track 
is  laid  with  30  Ib.  rail  to  the  first  gathering  point  in  the  mine, 


Total  lengfh  of  hauJ  in  No.!,  heading  -4,400rt. 
Average  grade  in  favor  of  load  2°' 
Average  grade,  again  empties 
approx.  2.2% 

l"< 


CO.  HOUSES 

Kails 

Main  haulage  SO  Ib.  per  yard 
Room  and  entries  20  Ib.  per  yard 


=Side  Tracks, 
•Main  Haulage 


FIG.  23. — Haulage  layout  at  Shade  Coal  Co.'s  Mine  in  Pennsylvania. 

beyond  which  20  Ib.  rail  are  used.  A  map  of  the  haulage 
arrangement  is  shown  in  the  accompanying  drawing,  Fig.  23. 

The  first  motor  was  installed  in  October,  1911,  and  was  a 
7-ton  machine,  with  a  4-cylinder,  4-cycle  engine,  having  a  6  in. 
bore  and  stroke  and  developing  35-hp.  at  800  r.p.m.  The  second 
motor  was  a  duplicate  of  the  first  and  was  put  in  service  in 
March,  1913. 

The  coal  at  this  mine  is  hauled  in  one  hundred  and  eighty- 
five  1000-lb.  mine  cars,  having  a  capacity  of  2500  Ib.,  all  of 
these  cars  being  equipped  with  plain  bearings.  The  outside 
and  main-entry  track  consists  of  30-lb.  rail,  and  back  of  the 
first  gathering  point  in  the  mine,  the  rail  is  20  Ib.  The  mini- 
mum radius  of  the  curves  is  approximately  30  ft. 


HAULAGE  COSTS  353 

The  first  gasoline  machine  was  put  in  operation,  October, 
1911,  and  was  7-ton,  36-in.  gage  locomotive.  It  has  a  vertical, 
4-cylinder,  4-cycle  engine,  6-in.  bore  and  6-in.  stroke,  which 
develops  35  hp.  at  800  r.p.m. 

Mule  haulage  was  used  before  the  gasoline  locomotives  were 
introduced  and  at  that  time  the  following  conditions  pre- 
vailed : 


OPERATING  CONDITIONS 

Length  of  haul  one  way,  2640  ft. 
Maximum  grade  in  favor  of  loads,  2|  per  cent 
Maximum  grade  against  empties,  5  per  cent 
Tonnage  per  month  of  24  working  days,  9600  tons 


COST  OF  MULE  HAULAGE 

10  mules  to  handle  tonnage  @  60c.  per  day $  6.00 

Above  day  rate  includes  feed,  harness  and  shoeing 
expenses 

Eight  drivers  @  $2 . 25  per  day 18 . 00 

Investment  in  10  mules  @  $200  is  $2000.  Assuming 
the  average  life  of  a  mule  is  5  years,  this  gives  a  20 
per  cent  depreciation  per  year,  24  working  days  per 
month,  the  cost  for  depreciation  per  day  will  be ....  1 . 38 

Interest  on  investment  at  6  per  cent  per  annum  per  day      0 . 41 

Mule  haulage  cost  per  day $25 . 79 

Total  mule  haulage  cost  per  month  of  24  working  days  618 . 96 

Mule  haulage  cost,  per  ton 0 . 064 

Mule  haulage  cost  per  ton-mile  traveled  by  loads ....       0 . 128 


GATHERING  COST  USING  MULES  AT  MINE  No.  1 

Tonnage  per  month  of  24  working  days,  9000  tons 

3  mules  for  gathering  @  60c.  per  day $1 . 80 

Three  drivers  @  $2.50  per  day 7. 50 

Mule  depreciation 0 . 417 

Interest  on  investment  at  6  per  cent  per  annum,  per 
day 0.125 

Total  gathering  expense  per  day $9 . 842 

Total  gathering  expense  per  month $236 . 208 

Gathering  expense  per  ton 0 . 0262 


354  COAL  MINING  COSTS 


MAIN  ENTRY  GASOLINE  HAULAGE  AT  MINE  No.  1 

Length  of  haul  one  way,  4400  ft. 
Maximum  grade  in  favor  of  loads,  6  per  cent 
Average  grade  in  favor  of  loads,  2  per  cent 
Maximum  grade  against  empties,  5  per  cent. 
Average  grade  against  empties,  2.2  per  cent 
Tonnage  per  month  of  24  working  days,  9000  tons 

Motorman  for  24  days  @  $2 . 75 $66 . 00 

Trip  rider  for  24  days  @  $2.75 66.00 


Total  labor $132.00 

384  gal.  of  gasoline  per  month  @  15£c .  .  .   $59 . 52 

24  gal.  engine  oil  per  month  @  30c 7 . 20 

6  gal.  black  oil  per  month  @  15c 0 . 90 

6  Ib.  cup  grease  per  month  @  15c 0 . 90 

Waste  per  month 1 . 00 

Total  supplies 69.52 

Repairs: 

Material $28.95 

Labor..  6.86 


Total  repairs 35 . 81 

Depreciation  on  locomotive  at  10  per  cent  per 

annum,  per  month 29 . 16 

Interest  on  locomotive  investment  at  6  per  cent  per 

annum,  per  month 17 . 50 


Total  operating  cost  per  month $283 . 99 

Operating  cost  per  day $11 . 83 

Operating  cost  per  ton 0.0315 

Operating  cost  per  ton-mile  traveled  by 
loads..  0.0379 


GATHERING  COST  USING  MULES  AT  MINE  No.  3 

Tonnage  per  month  of  24  working  days,  3,500  tons 

1  mule  required  for  gathering  at  60c.  per  day $0.60 

1  driver  required  per  day 2 . 40 

Mule  depreciation  per  day 0 . 138 

Interest  on  investment  at  6  per  cent  per  annum,  per 

day 0.0416 


Total  gathering  expense  per  day $3 . 1796 

Gathering  expense  per  month $76 . 308 

Gathering  expense  per  ton 0.0218 


HAULAGE  COSTS  355 

MAIN  ENTRY  GASOLINE  HAULAGE  AT  MINE  No.  3 
Length  of  haul  one  way,  3100  ft. 

Motorman  for  24  days  at  $3  per  day $72 . 00 

Trip  rider  for  24  days  at  $2 . 75  per  day 66 . 00 


Labor $138.00 

120  gal.  gasoline  per  month  @  15£c $18.60 

4  gal.  engine  oil  per  month  @  30c 1 .20 

2  gal.  black  oil  per  month  @  15c 0 . 30 

2  Ib.  cup  grease  per  month  @  15c 0.30 

2  Ib.  waste  per  month  @  15c 0.30 


Total  supplies 20.70 

Repairs : 

Material $28.95 

Labor..  6.86 


Total  repairs 35 . 81 

Depreciation  on  locomotive  @  10  per  cent  per 

annum,  per  month 29 . 16 

Interest  on  locomotive  investment  @  6  per  cent  per 

annum,  per  month , ? 17 . 50 


Total  operating  cost  per  month $241 . 17 

Operating  cost  per  day $10.04 

Operating  cost  per  ton 0.0691 

Operating  cost  per  ton-mile 0. 1173 

Mine  No.  3  is  under  development,  and  the  motor,  as  well 
as  the  gathering  mule  and  driver,  are  not  working  to  exceed 
40  per  cent  of  the  time.  The  gasoline  consumption,  and  the 
amount  of  sand  used,  varies  with  the  weather.  The  gasoline 
consumption  runs  from  4  to  7  gal.  per  day,  bad  weather  caus- 
ing more  wheel  slippage  and  a  higher  engine  speed,  and,  there- 
by, increasing  gasoline  fuel  consumption. 

SUMMARY  BASED  ON  EXPERIENCE  IN  MINE  No.  1 

Mule  haulage  cost  per  ton $0 . 064 

Gasoline  haulage  per  ton 0 . 0315 

Saving  per  ton 0. 0325 

Mule  haulage  cost  per  ton-mile 0 . 128 

Gasoline  haulage  per  ton-mile 0 . 0379 

Saving  per  ton-mile 0 . 0901 

The  following  is  a  typical  example  of  gasoline  consumption 
011  a  7-ton  motor,  developing  50  hp.  at  a  speed  of  500  r.p.m. 
On  a  break  test  this  motor  would  consume  about  a  pint  of 


356  COAL  MINING  COSTS 

gasoline  per  horsepower-hour,  equal  to  50  pints  or  6*4  gal.  per 
hour.  Theoretically,  therefore,  this  motor  would  consume  50 
gal.  of  gasoline  per  8-hr,  shift,  while  in  actual  practice  the 
consumption  varies  from  15  to  18  gal.  indicating  that  the  aver- 
age horsepower  developed  varies  between  the  same  figures. 

The  following  are  some  of  the  results  obtained  with  an 
Otto  internal  combustion  mine  locomotive,  working  at  the  Bar- 
ton mines  in  Nottingham,  England,  about  1910. 

The  length  of  haul  inside  the  mine  is  about  700  yd.  and 
on  the  surface  about  l1/^  miles.  With  the  previous  horse  trac- 
tion one  round  trip  on  the  surface  line  occupied  over  two  hours, 
including  shunting  at  the  tipple,  hauling  a  train  of  10  loaded 
wagons  of  about  20  tons  total  gross  load.  The  locomotive 
makes  one  trip  with  the  same  number  of  wagons  in  three- 
fourths  of  an  hour  regularly,  and  in  some  cases  in  35  min. 
The  line  is  for  the  greater  part  level  but  partly  in  favor  of 
the  loads.  The  heaviest  grades  are  0.77  per  cent  against  loads 
and  4  per  cent  against  empties. 

The  gasoline  consumption  during  27  working  days  was  38 
gal.  which  equals  1.4  gal.  per  day  from  7  a.m.,  to  5  p.m.  Dur- 
ing this  time  1274  net  tons  of  stone  were  hauled.  This  shows 
that  with  one  gallon  of  gasoline,  33%  net  tons  were  covered; 
at  a  tare  of  10  cwt.  per  wagon  this  represents  a  total  gross 
load  of  47.6  tons,  this  being  over  a  line  of  1%  miles,  so  that 
71.4  ton-miles  were  covered  with  one  gallon  of  gasoline.  The 
locomotive  is  fitted  with  two  speed-gears  in  either  direction, 
i.e.,  for  3%  and  7%  miles  per  hour. 

This  locomotive  has  replaced  six  horses  which  cost  $14.40 
per  week  in  fodder  alone,  while  also  requiring  four  boys.  The 
locomotive  only  requires  one  driver  and  one  boy  for  shutting 
the  gates,  when  crossing  the  roads.  The  gasoline  consumption 
per  week  of  about  Sy2  gal.  at  16c.  exclusive  of  rebate,  amounts 
to  about  $1.34  per  week.  . 

The  Germans  and  Austrians  were  the  pioneers  in  the  use 
of  gasoline  motors  for  mine  use.  In  1910  it  was  estimated  that 
there  were  about  300  of  these  in  use  in  various  parts  of  the 
world. 

Gasoline  motor  haulage  costs  were  estimated  in  1910  on 
underground  haulage  work  where  the  tracks  were  inferior  and 
curves  sharp  at  2.4  to  2.6c.  per  ton-mile.  Actual  working  costs 


HAULAGE  COSTS 


357 


taken  over  a  sufficiently  long  time  to  give  reliable  results  were 
found  on  the  surface  track  to  be  as  low  as  1.2c.  per  ton-mile 
after  deducting  20  per  cent  for  amortization. 

Fuel  consumption  on  one  type  of  gasoline  motor  was  found 
to  be  slightly  less  than  0.1  gal.  per  hp.-hr.  when  working  under 
full  load.  As,  however,  the  motor  is  never  run  at  full  load 
except  when  starting  and  running  up  grade,  it  is  found  that 
0.05  gal.  per  hp.-hr.  is  the  normal  consumption. 

Storage  battery  and  trolley  motor  haulage  costs. — Some 
excellent  figures  on  repairs  and  maintenance  costs  of  storage 
battery  motors  on  metal  mine  work  at  the  Bunker  Hill  and 
Sullivan  Mine  are  given  on  p.  229,  Vol.  51  A.I.M.M.E.  It  was 
found  there  that  low  voltage,  as  compared  to  the  500-volt  d.c. 
used  on  the  trolley-type  locomotive,  practically  eliminates 
brush  and  commutator  troubles,  which  always  have  been  a 
source  of  heavy  expense.  About  the  only  charge  against  the 
batteries  is  the  time  of  one  man  for  a  few  minutes  each  morn- 
ing, giving  them  the  daily  inspection  and  refilling  the  cells 
with  distilled  water  to  replace  that  evaporated  during  the 
previous  day — the  amount  of  distilled  water  required  for  three 
batteries  being  about  20  gal.  per  week.  In  addition,  there  is 
a  monthly  charge,  not  exceeding  $10  per  battery,  for  cleaning 
and  overhauling.  The  principal  source  of  repair  expense  on 
the  locomotives  was  for  new  wheels. 

The  figures  given  below  are  the  total  average  monthly  cost 
of  repair  and  upkeep  from  the  date  of  installation  to  Novem- 
ber 1,  1914,  and  also  include  the  cost  of  installation,  which 
was  quite  large,  because  the  Battery  boxes  had  to  be  altered 
and  partly  rebuilt,  to  adapt  them  to  the  company's  charging 
system,  and  to  protect  them  from  the  water  issuing  from  the 
chutes  under  which  they  pass. 


• 

Average 

Average 

Repair 

Monthly 

Monthly 

Cost 

Repair 

Tonnage 

in 

Cost. 

Hauled. 

Cents. 

2|-ton  Jeffrey,  No.  11  level 

$48  513 

13,501 

0.359 

4-ton  Westinghouse,  No.  12  level  

55  456 

12,755 

0.434 

4-ton  Gen  Electric  No  13  level 

28  612 

2,406 

1  189 

358  COAL  MINING  COSTS 

The  last  figure  is  high  because  the  costs  are  figured  only 
on  the  tonnage  of  ore  hauled,  and  most  of  the  material  handled 
by  this  motor  was  waste. 

A  comparison  of  these  costs  with  the  trolley  motors  which 
they  replaced  is  interesting.  They  cover  a  period  of  two  months 
in  the  first  case  and  four  months  in  the  second,  so  that  the 
figures  must  not  be  taken  to  represent  an  average  cost  over 
a  long  period.  Separate  repair  costs  were  kept  for  all  motors 
until  January,  1913,  which  accounts  for  the  short  period  taken 
for  the  above  motors,  when  it  is  remembered  that  they  were 
replaced  by  storage-battery  motors  in  March  and  June  respec- 
tively, in  the  same  year.  The  2%-ton  Jeffrey  trolley  motor  on 
the  No.  11  level  had  an  average  monthly  repair  cost  of  $39.88 
as  compared  with  $93.912  for  the  4i/2-ton  General  Electric 
working  on  the  No.  12  level ;  the  average  monthly  tonnage 
handled  by  the  two  machines  was  9154  for  the  2y2-ton  and 
14,645  for  the  41^-ton  motors;  and  the  repair  costs  per  ton 
were  4.35c.  and  6.41c.  respectively. 

These  figures  do  not  include  the  initial  cost  of  upkeep  and 
the  trolley  wires  and  track  bonding,  which  kept  two  men  busy 
practically  all  the  time,  and  which  was  consequently  a  heavy 
expense.  No  separate  costs  were  kept  for  the  two  levels,  how- 
ever, so  they  may  be  omitted  in  this  connection.  It  has  been 
estimated  by  the  company's  electrical  engineers  that,  with  a 
few  minor  improvements  in  the  charging  system,  and  a  "better 
understanding  of,  and  more  careful  attention  given  to  the 
operation  of  these  motors  by  the  motormen,  the  cost  of  repairs 
and  operation  would  be  75  per  cent  less  than  the  trolley  motors 
doing  the  same  work. 

Repair  records  on  a  4-ton  Westinghouse  storage-battery 
motor  at  the  Big  Five  Tunnel  in  Colorado,  covering  a  6-months 
period,  showed  a  cost  of  $12.60  per  month,  or  0.8c.  per  ton- 
mile.  The  motor  was  doing  very  light  work  during  three 
months  of  this  time  so  that  the  repair  costs  may  be  relatively 
high ;  for  two  months  during  which  time  it  was  working  under 
a  more  nearly  full  load,  the  repair  cost  per  ton-mile  was  0.65c. 

One  of  the  chief  objections  advanced  to  the  storage  battery 
motor  is  the  cost  of  the  batteries  and  their  comparatively  short 
life  which  ranges  from  two  to  four  years.  It  is  doubtful,  how- 
ever, if  this  extra  charge  against  the  motor  will  exceed  the 


HAULAGE  COSTS  359 

cost  of  copper  wire,  bonding  and  twin  cables  and  cost  of 
upkeep  for  the  trolley  type  motor.  It  must  also  be  remem- 
bered that  the  storage  battery  motor  uses  less  than  half  the 
power  required  to  operate  the  cable  and  reel  motor  which  it 
usually  replaces  on  secondary  haulage. 

Constant  fluctuations  in  output  in  different  sections  of  the 
mine  makes  it  difficult  to  operate  trips  on  any  regular  time- 
table as  in  the  case  of  railroads.  This  is  due  largely  to  the 
fact  that  there  are  so  many  elements  in  the  movement  of  the 
cars,  a  delay  or  accident  in  any  one  of  which  would  derange 
the  whole  schedule.  The  system  here  described  was  in  use  in 
a  mine  of  the  Sterling  Coal  Co.  in  Ohio. 

The  coal  in  this  mine  is  handled  over  one  main  entry  off 
which  there  are  11  butt  entries  having  from  12  to  15  working 
rooms  each.  The  motive  power  consists  of  one,  8-ton  trolley 
motor  on  the  main  haul,  two  S^-,  one  4-,  and  one  4%-ton 
motors  for  the  butt  entry  hauls  and  eleven,  2!/^-ton  storage 
battery  motors  for  gathering.  The  8-ton  motor  on  the  main 
haul  handles  36-car  trips  taking  all  the  loads  from,  and  placing 
all  the  empties  at  a  sidetrack  along  the  main  haul. 

The  butt-entry  locomotives  take  the  empties  from  the  side- 
track in  trains  of  12  cars  and  deliver  them  to  one  of  the  2y2- 
ton  storage-battery  locomotives.  Each  butt-entry  locomotive 
takes  care  of  three  butt  entries. 

When  the  butt-entry  locomotive  brings  in  a  trip  of  12 
empties,  this  trip  together  witK  the  21^-ton  storage-battery 
locomotive  is  pushed  into  a  room.  In  the  meantime  the  butt- 
entry  locomotive  makes  into  a  train  the  12  loaded  cars  which 
have  been  placed  on  the  entry  by  the  storage-battery  locomo- 
tive. After  the  loaded  trip  is  coupled  up  it  is  pulled  down  to 
the  room  where  the  empty  trip  and  storage-battery  locomotive 
are  waiting  and  coupled  onto  them.  The  empty  trip  is  then 
pulled  with  its  battery  locomotive  out  on  the  entry,  after  which 
the  empty  trip  is  uncoupled,  the  loaded  trip  taken  to  the  part- 
ing by  the  butt-entry  locomotive,  and  the  storage-battery  loco- 
motive then  proceeds  to  distribute  the  empty  cars  and  push 
them  to  the  face  of  the  rooms. 

The  reason  for  pushing  the  empty  trip  into  a  room  and 
pulling  it  out  again  with  the  trolley  locomotive  is  to  save  the 
storage  battery  from  handling  the  trip  of  empties  while  the 


360  COAL  MINING  COSTS 

trolley  locomotive  is  on  the  entry.  The  twelve  loads  together 
with  two  other  similar  trips  taken  from  two  other  butt  entries 
are  taken  to  the  siding  to  make  up  a  trip  of  36  loads  for  the 
main-haulage  locomotive. 

With  this  system  of  haulage,  a  schedule  is  maintained 
approximately  as  follows:  One  entry  produces  12  cars  per 
.hour.  This  means  that  the  21^-ton  storage  battery  locomotive 
in  each  hour  places  12  empties  at  the  face  and  takes  12  loads 
away  to  the  entry  leaving  them  just  outside  the  room  neck. 
When  these  loads  have  been  placed  on  the  entry,  the  butt-entry 
trolley  locomotive  comes  along,  leaves  12  empties  and  picks 
up  the  trip  of  12  loads,  taking  it  to  the  sidetrack. 

This  trolley  locomotive  comes  into  each  butt  entry  once 
every  hour,  and  inasmuch  as  it  has  three  entries  to  take  care 
of,  there  are  20  min.  available  for  taking  care  of  each  entry. 
The  main-haulage  locomotive,  handling  36  cars  per  trip,  must 
make  four  round  trips  per  hour,  or  a  round  trip  every  15  min. 

The  average  weight  of  coal  loaded  into  each  car  is  2700  Ib. 
and  at  a  rate  of  production  of  12  cars  per  hour  per  butt  entry 
for  11  butt  entries,  the  capacity  of  this  haulage  system  is 
approximately  178  tons  per  hour.  On  this  basis,  the  haulage 
capacity  per  day  of  8  hr.,  is  1424  tons. 

Prior  to  the  installation  of  these  locomotives  the  cars  were 
handled  in  rooms  by  pushers,  three  being  required  on  each 
butt  entry.  Now  the  motorman  on  the  locomotive  is  the  only 
man  required  to  handle  these  cars.  It  is  estimated  that  the 
saving  thus  effected  will  amount  to  approximately  $15,000 
per  year. 

A  comparison  between  this  three-element  and  a  two-element 
haulage  system  which  might  be  used  in  its  place,  may  be  of 
interest.  The  capacity  of  each  storage-battery  locomotive  in 
an  8-hr,  day  under  the  present  system  is  96  cars.  If  these  loco- 
motives were  eliminated  and  the  butt-entry  locomotives  pro- 
vided with  cable  reels,  to  enable  them  to  go  up  into  the  rooms 
to  place  the  empties  and  pull  the  loads,  a  cable-reel  locomotive 
would  be  required  on  each  entry.  The  cost  of  a  storage-bat- 
tery locomotive,  such  as  is  used  here,  and  that  of  a  cable-reel 
locomotive,  such  as  would  be  required,  are  practically  the 
same. 

With  the  two-element  system,  then,   only  11   locomotives 


HAULAGE  COSTS  361 

would  be  required  as  compared  with  15  under  the  present 
system,  saving  the  installation  of  four  locomotives  representing 
an  investment  of  approximately  $7200.  Assuming  depreciation 
and  interest  at  25  per  cent,  this  investment  costs  approximately 
$1800  per  year.  On  the  other  hand,  with  cable-reel  locomotives, 
two  men  would  be  required  on  each  locomotive,  or  a  total  of 
22  men  for  11  locomotives. 

With  the  present  system,  only  one  man  is  required  on  each 
storage-battery  locomotive  and  two  men  on  each  butt-entry 
locomotive,  making  a  total  of  19  men.  At  $60  per  month  per 
man,  the  present  system  thus  effects  a  saving  of  $2160  per 
year,  which  more  than  offsets  the  interest  and  depreciation  on 
the  added  investment.  To  this  should  be  added  the  freedom 
from  cable  trouble  and  expense,  freedom  from  the  expense  of 
more  carefully  laid  and  bonded  track  in  rooms  and  the  possi- 
bility of  operating  on  a  smaller  generator  equipment.  Thus  the 
economic  advantages  of  this  three-element  system  become  evi- 
dent. No  doubt  there  are  many  operations  where  this  plan, 
which  has  proven  its  advantages  and  economy  at  this  mine, 
can  be  applied  with  equal  success. 

While  the  storage-battery  locomotive  is  of  course  not  abso- 
lutely safe  in  the  presence  of  gas,  the  sparks  it  makes  are  not 
near  the  roof,  so  the  danger  is  lessened,  for  the  gas  must  be 
in  large  quantity  if  it  is  to  settle  so  low  as  to  be  ignited. 

Another  increase  of  safety  with  the  battery  locomotive 
results  from  its  shortened  electric  circuit.  The  current  does 
not  pass  from  the  power  house  to  the  motor  and  back  through 
the  rails  to  the  generator,  but  the  circuit  is  contained  within 
the  locomotive — not  even  the  wheels  are  included  in  its  range. 

Thus  if  it  is  true  that  there  is  a  risk  of  stray  currents  pre- 
maturely igniting  shots,  the  current  of  the  storage-battery  loco- 
motive can  meet  the  accusation  with  a  perfect  alibi.  And,  of 
course,  as  there  is  no  conducting  wire,  there  can  be  no  short 
circuits  to  ignite  gas,  coal  dust  or  wooden  structures. 

The  travel  of  the  electric  current  along  the  drawbars  and 
couplings  of  a  train  of  mine  cars  has  always  been  an  objection 
to  the  trolley  locomotive,  as  it  has  occasionally  shocked  men 
and  ignited  powder.  For  this  reason  operators  have  sought  to 
improve  the  grounding  of  the  traction  rigging  and  to  make 


362  COAL  MINING  COSTS 

powder  cars  relatively  nonconducting  throughout.  In  other 
cases  the  powder  car  has  been  hauled  around  by  a  mule. 

The  battery  locomotive,  however,  having  its  current  self- 
contained,  does  not  offer  any  such  risk.  It  seems  that  it  would 
serve  admirably  for  hauling  men  and  transporting  material, 
explosive  and  otherwise,  into  and  out  of  the  mines.  By  switch- 
ing off  the  electric  current  on  the  main  haulage  road,  the  load- 
ing of  men  on  the  man  trip  and  their  passage  along  the  road 
at  the  beginning  and  end  of  the  day  would  be  robbed  of  its 
dangers,  some  of  which,  though  unnecessary,  are  unavoidable 
so  long  as  careless  and  ignorant  men  are  employed. 

The  use  of  a  section  insulator  at  the  point  of  embarkation 
or  disembarking  has  been  occasionally  adopted  and  has  prob- 
ably saved  many  lives. 

In  mines  where  all  or  part  of  the  night  load  is  quite  light, 
it  is  customary  to  delay  pumping  till  the  mines  are  closed  down. 
If  undercutting  is  done  at  night  and  the  coal  is  loaded  and 
hauled  out  by  day,  there  is  a  third  shift  during  which  the 
storage-battery  locomotives  can  be  charged.  This  tends  to 
keep  the  load  curve  even  and  to  save  in  expense.  Additional 
boosting  of  the  locomotive  batteries  can  be  performed  during 
the  lunch  hours  and  when  shifts  are  being  changed,  should 
the  batteries  need  it  and  should  the  men  walk  to  their  work. 

It  has  been  generally  thought  that  the  storage-battery 
repairs  would  be  excessive,  but  offsetting  this  there  are  no 
trolley  harps,  wheels  and  supports  to  be  maintained  in  con- 
dition. The  chance  of  the  armature  bearings  and  poles  heat- 
ing or  rubbing  is  about  the  same  in  both  battery  and  trolley 
locomotives.  The  battery  renewals  might,  however,  cost  so 
much  that  the  cost  of  harps,  wheels  and  supports  would  appear 
only  a  trifling  drawback  to  electric  locomotives. 

Compressed-air  and  electric  haulage  costs. — The  general 
superiority  of  electricity  for  underground  haulage  has  been  too 
well  established  by  its  widespread  popularity  during  the  last 
two  decades  to  admit  of  any  serious  controversy  as  to  the 
relative  economy  of  it  and  the  compressed-air  haulage 
methods.  Certainly  in  any  event,  the  economic  possibilities 
of  the  air-motor  are  limited  to  specific  and  unusual  conditions 
such  in  gaseous  mines  where  there  would  be  danger  in  the  use 
of  the  electric  motor  and  even  here  the  motor  would  find  its 


HAULAGE  COSTS 


363 


greatest  application  for  secondary  haulage  or  gathering  pur- 
poses. So  long  as  it  is  still  used,  however,  a  few  examples  of 
its  costs  of  installation  and  operation  will  be  of  value. 

The  accompanying  table  excerpted  from  Vol.  34,  A.I.M.M.E., 
gives  the  comparative  cost  of  electric  and  compressed-air  haul- 
age as  worked  up  by  the  H.  K.  Porter  Co.,  covering  results 
with  the  fourth  and  fifth  motors  built  by  that  concern.  They 
represent  the  results  of  operations  at  the  Glen  Lyon  plant  in 
1898.  The  table  gives  the  cost  per  ton  in  preference  to  the 
ton-mile  basis,  because  the  delays  at  terminals  forms  so  large 
a  part  of  the  time  lost  that  it  remains  a  fixed  quantity,  regard- 
less of  the  length  of  haul  so  that  the  best  showing  on  a  ton- 
mile  basis  is  only  obtained  on  long  hauls. 


Com- 
pressed 
Air 

ELECTRICITY 

Estimated 

Actual 

Number  of  working  days  during  year 

160 

*2362£ 

$1.16 
4.20 
3.20 

200 
989 

$1.20 
4.23 
3.20 
1.67 
5.95 

14H 
989 

$2.84 
9.31 
3.61 
3.68 
8.42 
0.46 
0.61 
2.50 
J8.17 
4.41 

0.35 
0.74 

Output  per  day  in  tons  

Cost  per  day: 
Engineer,  powerhouse  

Motorman 

Helpers  (brakeman)  

Electrician  

Repairs  to  motors 

0  74 

Repairs  to  line  

Repairs  to  generator         

0.57 

Fireman  • 

Depreciation  

14.74 
4.73 

1.63 
0.25 
0.47 
2.32 

J5.20 

Interest  ...    . 

Interest,  repairs  and  depreciation,  174  hp. 
boiler  

0.22 

Oil  and  waste  for  motor 

Oil  and  waste  for  generator  

Steam  (fuel  and  firing)  

Totals 

$24.01 
0.01015 

$21.67 
0.02192 

$45  .  10 
0.0456 

Cost  per  ton  

*  Tons  of  coal  hauled, 
t  At  5  per  cent, 
t  At  3  per  cent. 


364  COAL  MINING  COSTS 

In  this  table  interest  on  the  compressed-air  motors  is  cal- 
culated at  5  per  cent  and  at  3  per  cent  on  the  electric  motors. 
The  column  headed  "Actual"  gives  the  results  accomplished  at 
the  electric-haulage  plant  of  the  No.  2  Mine  of  the  Hillside 
Coal  &  Iron  Co.;  under  the  column  " Estimated"  those  results 
are  given  as  re-calculated  on  the  assumption  of  200  working 
days  in  the  year  instead  of  141.25. 

About  1900  one  of  the  larger  bituminous  coal  companies 
installed  a  compressed-air  haulage  system  consisting  of: 

1  compound-steam  three-stage  air  compressor,  800  Ib.  pres- 
sure. 

5300  ft.  5-in.,  triple  strength,  pipe-line. 

3600  ft.  2  in.,  triple  strength,  pipe-line. 

Three  14-ton  motors. 

The  cost  of  this  installation,  exclusive  of  boiler  plant  was 
approximately  $37,000.  The  time  consumed  for  each  trip  was 
45  min.  There  were  two  charging  stations  and  the  compres- 
sor was  intended  for  the  operation  of  all  three  motors  but 
under  actual  operating  conditions  there  was  very  little  margin 
when  one  motor  was  in  operation.  The  cost  of  maintenance  of 
the  high-pressure  plant  and  the  motors  was  $8  and  $7.50, 
respectively,  per  day. 

An  electric-haulage  system  was  later  installed  at  this  mine, 
consisting  of  two,  150-kw.  generators,  directly  connected  to 
two  22  X  20  in.  simple  engines  and  six,  13-ton  electric  mine 
motors,  feeder  and  trolley  lines,  etc.,  the  entire  cost  of  which 
was  about  $42,000,  exclusive  of  the  boiler  plant.  One  of  these 
motors  performs  all  the  work  of  the  above  described  com- 
pressed-air plant,  making  the  round  trip  in  30  min.  and  in  addi- 
tion is  used  on  other  entries  for  two  or  three  hours  per  day. 
That  part  of  the  cost  of  the  electric  plant  properly  chargeable, 
for  the  purpose  of  comparison,  against  the  old  compressed-air 
haulage  plant  would  be  about  $7000.  Also  a  few  months  before 
the  compressed-air  plant  was  abandoned,  one  of  the  motors 
was  overhauled  at  an  expense  of  $2000. 


SECTION   IV 
TIMBERING  COSTS 

The  enormous  quantity  of  timber,  poles,  lagging,  mine  ties, 
plank  and  lumber  in  general  used  in  mine  operations,  makes 
this  subject  of  great  importance,  and  upon  the  intelligent 
handling  of  this  material  in  the  future  will  depend  in  a  great- 
measure  the  cost  per  ton  output  of  coal. 

Although  wood  has  been  in  universal  use  since  creation, 
there  is  a  remarkable  lack  of  knowledge  as  to  its  structure 
and  behavior  in  its  various  uses,  by  those  who  might  be  ex- 
pected to  know  its  properties,  thereby  using  species  totally 
unfit  for  certain  purposes,  and  consequently  expensive  to  the 
company.  In  the  past,  when  timber  was  plentiful  and  cheap, 
it  mattered  little  in  many  cases  how  long  it  lasted,  as  the  ser- 
vice it  gave  was  ample  return  on  the  cost;  smaller  quantities 
being  consumed  then,  few  thought  it  necessary  to  study  the 
structure  and  behavior  of  wood  in  order  to  lengthen  its  service. 

How  to  buy  timber. — The  cost  of  wood  has  increased 
enormously  in  the  past  decade,  while  the  quality  is  steadily 
decreasing.  The  greater  care  in  its  selection  and  use  is  there- 
fore self-evident,  in  order  to  lengthen  its  life  by  elimination  of 
infected,  or  poor  quality.  The  same  care  should  be  used  in 
the  selection  of  the  proper  species  for  the  various  uses. 

It  is  necessary,  therefore,  to  use  improved  methods  in  select- 
ing prop  timber,  mine  plank  and  various  other  wood,  first,  by 
determining  the  species  meeting  the  most  important  require- 
ments, or  several  qualities  in  combination  as  shown  by  actual 
experience,  and  tests.  Second,  the  methods  of  procuring  the 
supply.  Third,  the  handling  and  preparation,  and  finally  the 
placing  in  the  mines. 

Conceding  that  the  present  prop  material  in  healthy  con- 
dition (such  species  as  Southern  short  leaf,  loblolly  pine  and 
spruce  pine)  is  probably  as  good  as  can  be  secured  at  a  reason- 

365 


366  COAL  MINING  COSTS 

able  first  cost,  the  next  question  is,  when,  how  and  where  to 
purchase  same.  Timber  cut  between  September  and  March  is 
preferable  to  timber  cut  during  the  spring  and  summer  months. 
Timber  that  is  cut  during  hot  weather  is  subjected  to  attack 
through  a  period  when  all  insect  life,  such  as  worms,  wood 
borers,  beetles  and  other  low  plant  life  (fungus  parasites)  are 
most  active  in  their  work  of  wood  destruction,  and  the  timber 
during  this  period  is  in  its  most  favorable  condition,  with  fresh 
green  sap,  inviting  attack  of  fungi  spores  and  borers  of  all 
kinds,  thereby  causing  the  first  signs  of  decay.  Experience  has 
shown  that  summer-cut  timber  does  not  give  the  satisfaction, 
nor  will  it  give  as  long  service  as  winter-cut  stock. 

Black  oak  and  red  oak  are  approximately  of  equal  value 
with  short-leaf,  loblolly  and  spruce  pine,  all  of  these  being  an 
easy  prey  of  fungi,  when  in  contact  with  soil,  while  white  oak 
is  of  greater  strength  and  durability.  A  white-oak  plank,  one 
inch  full  thickness,  is  equal  to  an  inch  and  a  half  pine  plank 
of  equal  width  at  an  approximate  decrease  of  one-third  in  cost 
per  surface  foot. 

Because  of  the  use  of  heavy  electric  motors  instead  of  mule 
haulage  in  mines,  it  is  essential  that  improved  provision  be 
made  to  take  care  of  mine  tracks,  by  the  purchase  of  mine 
ties,  manufactured  to  a  rigid  specification,  from  live  winter- 
cut  white  oak,  chestnut  oak  and  young  chestnut  properly 
seasoned  before  using,  thereby  reducing  purchases  and  track 
repairs. 

Sound  pine  trees  cut  down  in  the  winter  season  and  cut 
into  log  lengths,  stripped  of  their  bark  and  piled  in  layers 
with  sticks  between  each  layer  so  that  a  free  circulation  of 
air  can  pass  through  the  pile,  will  harden  the  exterior  juices. 
This  will  form  a  coating  which  will,  to  a  great  extent,  furnish 
protection  from  exterior  checking  and  materially  resist  the 
attack  of  the  fungus,  spores  and  wood-destroying  insects,  which 
protection  it  does  not  have  if  cut  during  the  summer  season. 

Green  and  unpeeled  pine  timber  placed  in  mines  for  gang- 
way use  is  sure  to  give  short  service  and  minimum  strength. 
Consequently  such  timber  is  the  most  expensive  for  the  service 
it  renders.  In  such  condition  crushing  is  most  liable  and  decay 
sets  in  quickly. 

Tests  should  be  made  to  determine  the  most  efficient  species 


TIMBERING  COSTS  367 

for  each  particular  use,  bearing  in  mind  cost  and  service  and 
specifications  covering  such  needs  should  be  prepared  and 
purchases  and  inspection  made  accordingly. 

Timber  used  and  costs  for  United  States. — The  following 
statistics  on  the  timber  used  in  the  mines  of  the  United  States 
in  1905  are  based  upon  data  gathered  by  the  Forest  Service  in 
cooperation  with  the  United  States  Geological  Survey.  Nearly 
14,000  mines  were  selected  in  which  the  use  of  timber  seemed 
certain  or  possible,  and  from  more  than  5000  of  these,  reports 
were  received  showing  that  timber  had  been  used. 

It  will  be  noted  that  2940  bituminous  coal  mines  used 
nearly  $6,400,000  worth  of  timber,  while  216  anthracite  minetf 
used  over  $4,400,000  worth,  the  average  cost  of  timber  per 
mine  being  $2170  for  the  bituminous  and  $20,524  for  the 
anthracite  mines. 

It  will  also  be  noted  that  the  timber  used  in  the  216  anthra- 
cite mines  was  of  slightly  greater  value  than  that  used  in  1718 
mines  for  precious  metals.  The  much  higher  cost  of  the  tim- 
bering required  for  the  anthracite  mines  is  due  to  several 
causes.  In  the  first  place,  many  of  the  anthracite  workings 
lie  at  great  depths,  and  some  of  the  larger  properties  have 
many  miles  of  gangways  which  have  to  be  carefully  main- 
tained. They  are  below  water  level,  and,  as  a  result  of  the 
combined  action  of  air  and  mine  water,  the  timbers  decay 
rapidly.  Some  of  the  beds  are  of  enormous  thickness,  and 
require  vast  quantities  of  timber  in  the  construction  of  * l  square 
sets"  to  support  the  roof  and  preserve  the  workings  in  over- 
lying coal.  Moreover,  since  the  hills  in  the  immediate  vicinity 
of  the  anthracite  mines  have  been  largely  denuded  of  timber 
suitable  for  mine  supports,  operators  are  obliged  to  obtain 
their  supplies  from  considerable  distances. 

Pennsylvania,  with  524  mines,  used  37,826,000  cu.  ft.  of 
round  timber  and  55,716,000  board  feet  of  sawed  timber,  cost- 
ing altogether  $2,290,053.  The  average  cost  of  the  round  tim- 
ber was  3.5c.  per  cubic  foot  and  that  of  the  sawed  timber 
$17.39  per  thousand  board  feet.  In  Illinois  400  mines  used 
10,342,300  cubic  feet  of  round  timber,  costing  6c.  per  cubic 
foot,  and  7,025,000  board  feet  of  sawed  timber,  costing  $22.04 
per  thousand  board  feet,  the  total  cost  being  $778,186.  The 
325  mines  in  West  Virginia  used  6,716,000  cu.  ft.  of  round 


368  COAL  MINING  COSTS 

timber  and  19,645,000  board  feet  of  sawed  timber.  The  total 
cost  was  $561,061;  that  of  the  round  timber  being  4.6c.  per 
cubic  foot,  and  that  of  the  sawed  timber  $12.76  per  thousand 
board  feet.  Next  in  order  of  total  outlay  for  timber  is  Ohio, 
with  $471,730 ;  Iowa  with  $232,148 ;  Indiana  with  $220,209 ;  and 
Alabama  with  $216,221.  None  of  the  other  states  used  over 
$200,000  worth  of  timber. 

Timber  is  more  expensive  in  Colorado  than  in  any  other 
state.  The  average  cost  of  the  round  timber  in  that  state  was 
11. 6c.  per  cubic  foot  and  that  of  the  sawed  timber  $33.76  per 
thousand  board  feet.  The  large  amount  of  mining  has  made 
a  heavy  demand  for  timber,  and,  although  most  of  the  round 
timber  is  obtained  locally,  much  of  the  sawed  timber  must  be 
shipped  in  from  considerable  distances  at  high  freight  rates. 
Round  timber  in  Wyoming  and  New  Mexico  cost  10.4  and 
10.5c.  per  cubic  foot,  respectively,  or  nearly  as  much  as  in 
Colorado ;  but  the  fact  that  sawed  timber  was  obtained  locally 
kept  its  price  down  to  $16.93  per  thousand  in  Wyoming  and 
$12.18  in  New  Mexico.  The  lowest  average  price  reported  for 
round  timber  was  3.3c.  per  cubic  foot  in  Indiana,  and  for  sawed 
timber  $5.58  per  thousand  in  Washington.  It  must  be  borne 
in  mind,  however,  that  in  many  cases  the  timber  used  was  cut 
from  land  belonging  to  the  mine  operators,  and  the  cost  in- 
cludes only  cutting  and  hauling. 

Reports  were  received  from  216  collieries  in  the  anthracite 
regions  producing  approximately  83  per  cent  of  the  total 
anthracite  tonnage  of  the  United  States.  Figures  for  the 
remaining  17  per  cent  were  computed,  using  as  a  basis  the 
reports  actually  received,  assuming  that  conditions  and  require- 
ments were  uniform  throughout  the  state.  The  results  of  the 
tabulation  show  that  121,565,000  ft.  board  measure  of  sawed 
timber  (equivalent  to  10,130,000  cu.  ft.)  and  52,440,000  cu.  ft. 
of  round  timber  were  used  during  1905. 

The  total  value  of  the  sawed  timber  was  $1,842,000,  or  $15 
per  thousand  feet  board  measure.  The  total  value  of  the  round 
timber  was  nearly  double  that  of  the  sawed  timber,  being 
$3,468,000,  or  $6.60  per  100  solid  cubic  feet— the  approximate 
equivalent  of  the  average .  standard  cord  of  128  cu.  ft.  The 
total  value  of  the  round  and  sawed  timber  combined  was 
$5,310,000,  or  about  Sy2c.  per  long  ton  of  coal,  using  as  a  basis 


TIMBERING  COSTS 


369 


for  the  calculation  the  production  in  1905 — in  round  numbers 
61,000,000  long  tons. 

So  far  as  reported,  the  kinds  of  wood  have  been  tabulated 
separately,  but  in  many  cases  the  operators  were  unable  to 
furnish  information  in  regard  to  the  quantity  of  each  species 
used,  and  it  has  therefore  been  necessary  to  classify  a  large 
amount  as  "  mixed"  or  * 'miscellaneous." 


ROUND  TIMBER 

SAWED  TIMBER 

Kind 

Cubic  Feet 

Kind 

Board  Feet  ' 

Yellow  pine  

9,250,000 
6,220,000 
1,180,000 
590,000 
444,000 
236,000 
165,000 
115,000 
10,263,000 
477,000 
23,500,000 

Hemlock  

63,600,000 
14,200,000 
2,860,000 
1,740,000 
371,000 
328,000 
84,000 
28,642,000 
1,370,000 
8,370,000 

Oak            

Yellow  pine  

Hemlock 

Oak 

Pitch  pine  

Maple  

Chestnut     

Spruce       

Beech 

White  pine 

Jack  pine  

Pitch  pine  

Spruce                     .    . 

Mixed  hardwoods  .... 
Mixed  softwoods  
Miscellaneous 

Mixed  hardwoods  

Mixed  softwoods  
IVliscellaneous 

Total 

Total           

52,440,000 

.121,565,000 

Of  the  species  used  for  round  timber,  yellow  pine,  of  which 
a  large  amount  is  loblolly  pine  from  the  South,  furnishes  one- 
half.  Oak  ranks  next,  but  furnishes  a  much  smaller  propor- 
tion, according  to  the  reports.  The  proportion  of  oak  would 
unquestionably  be  increased  if  the  large  items  reported  as 
11  mixed  hardwoods"  and  "miscellaneous"  could  be  separated 
into  species,  and  it  is  not  improbable  that  oak  would  then  dis- 
place yellow  pine  in  rank. 

For  sawed  timber  hemlock  holds  first  place  in  quantity, 
while  yellow  pine  ranks  next.  The  amount  of  oak  reported  is 
doubtless  too  small,  but  an  explanation  is  found  in  the  classifi- 
cation for  " mixed  hardwoods"  and  ''miscellaneous,"  which 
contains  over  37,000,000  ft.  board  measure,  of  which  probably 
a  large  amount  is  oak. 


370  COAL  MINING  COSTS 

Computing  size  of  timber.  —  There  seems  to  be  no  definite 
rules  by  which  the  size  of  mine-gangway  timbers  may  be  cal- 
culated. Experience  largely  governs  their  use  and,  as  both 
experience  and  judgment  differ  widely,  in  many  cases  the 
timbers  are  either  too  small  or  too  large  for  the  work  required 
of  them.  The  compressive  stress  of  timbers  —  wood  or  steel  — 
is  often  neglected,  resulting  in  economic  waste. 

Owing  to  the  adjustment  of  stresses  underground,  it  is 
impossible  to  calculate  the  loading  of  the  timbers  with  mathe- 
matical accuracy;  however,  rules  may  be  formulated  that  will 
give  approximate  results.  The  experienced  miner  knows  the 
size  of  collar  best  adapted  for  certain  conditions  in  the  gang- 
way, so  that  by  taking  an  average  of  the  strength  of  collars 
used  in  several  different  gangways  we  arrive  at  a  value  that 
may  be  used  in  calculating  other  gangway  timbers. 

Where  the  strata  are  horizontal  or  the  dip  is  light  the  loads 
are  applied  normal  to  the  collar,  and  produce  therein  bending 
stresses,  while  compressive  stresses  result  in  the  legs,  each  leg 
taking  one-half  of  the  total  load  on  the  collar. 

Where  the  dip  of  the  strata  is  great,  however,  legs  and 
collars  may  be  subject  to  both  bending  and  compressive  stresses 
at  the  same  time.  The  compressive  stresses  in  the  collars  may 
safely  be  ignored,  as  they  are  taken  care  of  in  the  average 
bending  load  that  a  collar  will  support,  but  the  legs  must  be 
calculated  for  both  bending  and  compression.  This  bending 
load  per  foot  of  length  will  be  assumed  as  being  equal  to  that 
applied  to  the  collar. 

When  the  diameter  and  length  of  an  existing  collar  are  known 
the  safe  load  per  foot  of  length  that  it  will  support  may  be  calcu- 
lated by  the  formula: 


(1) 


in  which 

w  =  safe  load  per  foot  of  length  of  span  (lb.); 
d  =  least  diameter  of  collar  (in.); 
1  =  length  of  clear  span  (ft.); 

/=safe  unit  fiber  stress  (1200  lb.  per  sq.  in.  of  section,  for 
long-leaf  yellow  pine  and  white  oak,  and  900  lb.  per 
sq.  in.,  for  short-leaf  yellow  pine). 


TIMBERING  COSTS  371 

In  calculating  legs  for  compressive  stresses  only,  the  formula 

LJ700+15C+C2) 
Jc~    '    a(700+15c)     ' 

should  be  used,  in  which 

L  =  total  load  on  leg  (Ib.)  ; 
a  =  area  of  cross-section  of  leg  (sq.  in.)  ; 
c  =  length  of  leg  in  inches,  divided  by  its  least  diameter  in 
inches. 

The  diameter  of  the  leg  must  be  assumed  for  trial,  the  most 
economical  section  being  that  in  which  the  safe  unit  fiber  stress 
is  not  exceeded. 

The  following  practical  examples  will  show  the  correct  method 
of  using  these  formulas: 

Example  1. — What  size  legs  would  be  required  for  a  long-leaf 
yellow-pine  timber  set  in  a  gangway  12  ft.  wide,  with  clear  head- 
room of  8  ft.;  the  strata  being  horizontal?  Assume  2000  Ib.  per 
lineal  foot  for  load  on  collar. 

Solution—  2000X12  =  24,000  Ib.,  total  load  on  collar,  and 
24,000-^2  =  12,000  Ib.,  total  load  on  each  leg. 

Assuming  the  least  diameter  of  the  leg  to  be  4f  in.,  the  area 
of  section  is  0.7854  (4f)2=  15.03  sq.  in.;  and  c  =  8X!2^4f  =  21.94, 
which  substituted  in  Formula  2  gives  for  the  compressive  stress 
in  the  leg, 

12,000  (700+15X21.94+21.942) 
fc=          15.03(700+15X21.94)  l172  *'  per  sq'  m' 

Since  1172  Ib.  is  less  than  the  unit  stress  for  long-leaf  yellow 
pine  (1200  Ib.),  a  4f-in.  stick  will  support  the  load;  however,  in 
this  case,  it  would  be  advisable  to  use  a  larger  stick,  say  8-in. 
diameter,  to  allow  for  defects  in  the  wood  and  unforeseen  bending 
stresses. 

Example  2. — Design  a  long-leaf  yellow-pine  timber  set  for  a 
gangway  12  ft.  wide,  with  a  clear  headroom  of  8  ft.,  the  strata 
having  a  dip  of  45  deg.  It  has  been  found  that  2000  Ib.  per  lineal 
foot  of  collar  is  the  average  load  in  districts  where  the  dip  of  the 
strata  is  very  heavy;  so  we  will  assume  this  value  in  solving  the 
problem. 

Solution. — Reversing  Formula  1,  so  as  to  give  the  value  of  d 


372  COAL  MINING  COSTS 

and  substituting  the  values  given  for  the  load  per  lineal  foot  of 
collar,  unit  fiber  stress  and  length  of  span,  we  have 


wl2        3I  2000X122 

Say 


The  load  producing  compression  in  each  leg  is  24,000-^2  = 
12,000  Ib.  Assuming  as  before  the  same  bending  load  for  the  leg, 
as  for  the  collar  2000X8  =  16,000  Ib.  bending  load  on  each  leg. 
Then,  assuming  12^  in.  as  the  least  diameter  of  leg  and  solving 
for  the  allowable  unit  stress  due  to  compression,  we  have  by  For- 
mula 2,  since  c  =  8X!2-^12|  =  7.68. 

L_f  (700+  15c)  _  1200  (700+15X7.68)  _ 
a~700+15c+c2     700+15X7.68+7.682~  MH'Jju 

The  unit  stress  due  to  bending,  for  a  total  load  of  16,000  Ib., 
as  found  above,  is  calculated  by  solving  Formula  1  for  /;  thus, 

wl2  2000  X82 

m' 


The  unit  stress  due  to  compression  is 
L  12,000 


Jc 


0.7854  d2    0.7854  X12.52 


The  total  stress  in  the  leg  due  to  bending  and  compression  is 
therefore 

1008.2+97.8=  1106  Ib.  per  sq.  in. 

As  the  unit  stress  produced  by  the  loads  is  thus  shown  to  be 
nearly  equal  to  but  less  than  the  allowable  unit  stress,  a  12|-in. 
stick  is  the  most  economical  section  to  use. 

When  calculating  beams  of  special  cross-section,  as  steel 
I-beams,  it  is  customary  to  employ  an  expression  called  the 
"  section  modulus  "  of  the  beam.  The  section  modulus  (S)  of  a 
beam  is  a  value  that  multiplied  by  the  unit  fiber  stress  (/)  of  the 
material  gives  the  bending  moment  (M)  of  which  such  beam  is 
capable  of  supporting,  as  expressed  by  the  formula: 

M=fS  ......    .:  ..    .     (3) 

But,  for  a  beam  uniformly  loaded  and  supported  at  each  end, 
the  bending  moment  (M),  in  inch-pounds,  is 

,,     I2wl2     1  c     72  ,.N 

M=    g     =1.5wl2  .....     .     .     (4) 


TIMBERING  COSTS  373 

Combining  equations  3  and  4  gives  for  the  section  modulus 

5=1.5^  ........     (5) 

Example  3.  —  Design  a  steel  timber  set  for  a  gangway  12  ft. 
wide,  with  a  clear  headroom  of  8  ft.,  the  strata  having  a  heavy 
dip,  assuming  a  fiber  stress  for  steel,  /=  16,000  Ib.  per  sq.  in. 

Solution.  —  The  average  load  per  lineal  foot  on  the  collar  being 
assumed,  as  before,  w  =  2000  Ib.  and  the  span  being  1=12  ft.,  the 
required  section  modulus,  in  this  case,  is 

Kwl2      1.5X2000X122 

:L5T=     ~Woo~ 

Any  one  of  the  following  sections,  taken  from  steel  manufac- 
turers' handbooks,  will  be  satisfactory:  8-in.-32.5-lb.  Bethlehem 
girder  beam,  section  modulus,  28.6;  8-in.-32-lb.  Bethlehem  H-col- 
umn,  section  modulus  26.9;  10-in.-30-lb.  standard  I-beam,  section 
modulus,  26.8.  A  10-in.-30-lb.  I-beam  is  the  most  economical 
section,  but  an  H-beam,  or  girder  beam,  is  to  be  preferred  as  it 
has  a  broader  bearing  surface. 

Assuming  the  leg  to  be  a  10-in.-25-lb.  I-beam,  which  has  an 
area  a  =  7.37  sq.  in.;    radius  of  gyration,  r  =  0.97;    and  section 
modulus,  $  =  24.4;  using  Gordon's  formula  for  medium  steel,  fixed- 
end  column  and  solving  for  ultimate  strength  (P),  we  have 
50,000  50,000 


n 

+ 


36,000  r>          36,OOOX0.972 

39  308 
With  a  safety  factor  of  4,  the  safe  unit  stress  is     '       =  9827  Ib. 

Again,  assuming  the  same  unit  bending  load  for  the  leg  as  for 
the  collar,  w  =  2000  Ib.  per  lineal  foot,  and  the  length  of  the  leg 
being  I  =  8  f  t.  the  bending  moment,  applying  Formula  4,  is 
M=  1.5  wl2  =  1.5X2000X82  =  192,000  in.-lb. 

This  makes  the  unit  fiber  stress  due  to  bending 

,    M     192,000    7_.07, 

j  =  -«•  =    n..    =  7869  Ib.  per  sq.  in. 

The  unit  stress  due  to  compression,  for  a  12-ft.  span,  the  sec- 
tional area  of  the  leg  being  a  =  7.37  sq.  in.  is 
L     2000X12 


2X7.37 


374 


COAL  MINING  COSTS 


which  makes  the  total  unit  stress  due  to  bending  and  compression 
7869+1628  =  9497  Ib.  per  sq.  in. 

This  actual  stress  in  the  leg  is  less  than  the  safe  stress,  there- 
fore a  10-in.-25-lb.  I-beam  will  satisfy  the  conditions  assumed  in 
this  case. 

The  accompanying  table  sums  up  the  results  of  a  series  of 
tests  made  by  the  U.  S.  Forest  Service,  to  determine  the  effect 
of  knots  of  different  classifications  on  the  crushing  strength  of 
certain  varieties  of  timber.  It  will  be  noticed  that  in  some  cases 
the  presence  of  knots  seems  actually  to  increase  the  strength. 

RATIO  OF  RESULTS  OF  STRENGTH  TESTS  ON  KNOTTY  TIMBER  TO  RESULTS  ON 
CLEAR  TIMBER,  STRENGTH  OF  CLEAR  TIMBER  TAKEN  AS  UNITY 


Compressive 
Strength  at 
Elastic  Limit 
per  Square  Inch 

Crushing 
Strength  at 
Maximum 
Load 
per  Square  Inch 

Modulus  of 
Elasticity 
per  Square  Inch 

Douglas  fir: 
Pin  knots  . 

0  95 

0  94 

1  06 

Standard  knots  
Large  knots 

0.87 
0  78 

0.86 
0  78 

0.90 
0  71 

Western  larch: 
Pin  knots  
Standard  knots  .  .  . 

1.12 

0  98 

1.04 
0  89 

1.19 
1  00 

Large  knots  

0.98 

0.85 

Western  hemlock: 
Pin  knots     

0  96 

0  97 

1.00 

Standard  knots 

0  94 

0  91 

0  97 

Large  knots  

0.86 

0.83 

0.81 

Pin  knots  are  denned  as  sound  knots  %  in.  or  less  in 
diameter.  Standard  knots  are  defined  as  sound  knots  ranging 
from  1/2  to  11/2  in.  in  diameter.  Large  knots  are  also  sound 
knots  from  l1/^  in.  in  diameter,  up. 

The  accompanying  table  shows  what  different  sizes  of  steel 
rails  will  support  when  uniformly  loaded  and  from  these  loads, 
the  sizes  of  equivalent  standard  I-beams  and  the  different 


TIMBERING  COSTS 


375 


TABLE  COMPARING  STRENGTH  OF  STEEL  AND  WOOD  FOR  SUPPORTING  MINE 

ROOFS 

TEN-FOOT  SPAN 


Uni- 
form 
Load 
inLb. 

Re- 
quired 
Std. 
T-rail 

Standard 
I-beam 

In.    Lb. 

White  Oak 

Chestnut 

White  Pine 

Sawed 

Round 

Sawed 

Round 

Sawed 

Round 

1,015 

16 

3    16.5 

5X3 

5 

6X4 

5 

6X4 

5 

1,385 

20 

3    16.5 

5X4 

5£ 

6X5 

6| 

6X5 

ftj 

1,920 

25 

3    19.5 

6X4 

6| 

7X5 

71 

7X5 

71 

2,450 

30 

4    22.5 

6X5 

7* 

7X7 

8 

7X7 

8 

3,090 

35 

4    22.5 

7X5 

71 

8X7 

8| 

8X7 

8| 

3,840 

40 

5    29.3 

7X6 

8 

8X8 

9 

8X8 

9 

4,475 

45 

5    29.3 

7X7 

8| 

10X6 

9| 

10X6 

$i 

5,225 

50 

5    36.8 

8X6 

9 

10X7 

10 

10X7 

10 

6,290 

55 

6    36.8 

9X6 

9| 

10X8 

IQi 

10X8 

101 

7,140 

60 

6   36.8 

9X7 

10 

10X10 

11 

10X10 

11 

7,890 

65 

6   44.5 

10X6 

10 

12X7 

ill 

12X7 

iij 

8,955 

70 

7    45.0 

10X7 

«Hi 

12X8 

12 

12X8 

12 

9,700 

75 

7   45.0 

9X9 

11 

12X9 

12i 

12X9 

m 

10,765 

80 

7   45.0 

10X8 

11 

12X10 

13 

12X10 

13 

11,940 

85 

7    52.5 

10X9 

ni 

12X11 

13 

12X11 

13 

13,110 

90 

8    54.0 

10X10 

12 

12X12 

131 

12X12 

13| 

14,180 

95 

8    54.0 

12X8 

12 

12X13 

14 

12X13 

14 

15,670 

100 

8    60.8 

12X9 

12| 

12X14 

14| 

12X14 

141 

TWELVE-FOOT  SPAN 


845 

16 

3    16.5 

5X3 

5 

6X4 

6 

6X4 

6 

1,155 

20 

3    16.5 

5X5 

N 

6X5 

6| 

6X5 

6| 

1,600 

25 

4   22.5 

6X4 

« 

7X5 

7| 

7X5 

7* 

2,045 

30 

4   22.5 

6X5 

7 

7X7 

•    8 

7X7 

8 

2,580 

35 

5    29.3 

7X5 

7| 

8X6 

8| 

8X6 

81 

3,200 

40 

5    29.3 

7X6 

8 

8X8 

9 

8X8 

9 

3,735 

45 

6   36.8 

7X7 

81 
2 

10X6 

9| 

10X6 

9| 

4,355 

50 

6   36.8 

8X6 

9 

10X7 

10 

10X7 

10 

5,245 

55 

6   36.8 

8X8 

9| 

10X9 

11 

10X9 

11 

5,955 

60 

7    45.0 

9X7 

10 

11X8 

11 

11X8 

11 

6,580 

65 

7    45.0 

10X6 

10 

10X10 

111 

10X10 

111 

7,470 

70 

7    45.0 

10X7 

10| 

11X10 

12 

11X10 

12 

8,090 

75 

7    52.5 

10X8 

11 

12X9 

12| 

12X9 

12i 

8,980 

80 

8    54.0 

11X7 

11 

12X10 

13 

12X10 

13 

9,955 

85 

8   54.0 

10X9 

111 

12X11 

131 

12X11 

13| 

10,935 

90 

8    60.8 

10X10 

12 

12X12 

14 

12X12 

14 

11,825 

95 

9   63.0 

12X8 

12* 

13X11 

14 

13X11 

14 

13,070 

100 

9    63.0 

11X10 

12| 

13X12 

14| 

13X12 

14| 

NOTE. — Loads  given  in  table  are  the  safe  uniform  loads  that  T-rails  will  carry:  Other 
members  show  sizes  necessary  for  these  loads.  Timber  presumed  as  seasoned.  For  green 
timber  use  §  loads.  Factor  of  safety  6  (about).  Fiber  stress  white  oak,  1200  lb.;  white 
pine  and  chestnut,  800  lb.  Timber  to  be  placed  narrow  side  against  roof. 


376  COAL  MINING  COSTS 

wood  beams  have  been  calculated.  Thus,  taking  a  20-lb.  mine 
rail  it  is  seen  that  the  safe  uniform  load  it  will  carry  on  a 
10-ft.  span  is  1385  lb.,  then  following  this  line  to  the  right, 
under  the  same  conditions  it  is  seen  that  the  3-in.  16.5-lb. 
I-beam  will  do  the  same  and  save  3.5  lb.  per  yard,  or  about 
14  lb.  to  a  beam  of  12-ft.  length.  Under  white  oak  is  found  a 
5  X  4-in.  sawed  or  S^-in.  round  timber  for  this  load,  while 
chestnut  and  white  pine  require  a  6  X  5  in.  sawed  or  6^-in. 
round  timber. 

Quite  frequently  a  requisition  will  call  for  a  certain  sized 
timber,  say,  an  8-in.  round  timber,  to  be  14  ft.  long,  to  be  used 
for  a  12-ft.  span.  It  is  seen  by  the  table  that  a  4-in.  7.5-lb. 
I-beam  will  carry  the  same  load.  The  importance  of  the  place 
and  the  length  of  time  it  is  to  be  maintained  should  govern 
which  should  be  used.  Timber  in  a  mine,  if  it  carries  approxi- 
mately the  safe  load,  will  seldom  last  much  longer  than  18 
months,  and  the  replacement  plus  the  first  installation  will 
be  more  than  the  cost  of  placing  the  steel. 

This  table  is  presented  with  the  idea  that  it  will  be  used 
by  mine  officials  in  the  proper  selection  of  material  when  re- 
placing or  retimbering,  and  to  give  them  some  idea  what  their 
selection  of  sizes  will  carry  in  the  weight  of  roof  supported. 

The  Scranton  Mine  Cave  Commission  conducted  a  series  of 
interesting  experiments  to  determine  the  comparative  strength 
of  various  materials  used  for  supporting  mine  roofs,  a  summary 
of  the  results  of  which  are  given  in  the  accompanying  table. 

Timber  framing  equipment. — Where  a  large  amount  oi 
timber  framing  is  necessary,  say  sufficient  to  require  the  ser- 
vices of  four  to  five  framers  continuously,  the  introduction  of 
machinery  to  replace  these  men  should  be  considered.  This 
will  be  something  to  eliminate  the  use  of  the  two-man  cross- 
cut saw,  hewing,  gaging  and  framing  the  ends  as  required. 

A  regular  timber  framing  machine  for  this  work  will  cost 
from  $2500  to  $3000  (in  1914)  and  weigh  about  9000  lb.  It 
will  require  about  a  50-hp.  engine  and  boiler  to  drive  it  and 
will  then  only  cut  the  framing  gages  at  the  ends.  The  total 
cost  in  place  will  be  about  $7000  allowing  for  a  suitable  build- 
ing to  house  it. 

A  good  and  economical  substitute  for  this  is  a  slabber  and 
swinging  cut-off  saw  which  can  be  purchased  complete  for 


TIMBERING  COSTS 


377 


$420  (figures  as  of  1914).  This  equipment  can  be  erected  as 
shown  in  the  accompanying  illustration,  Fig.  1,  in  which  it 
will  be  noted  that  the  cut-off  saw  is  set  horizontal  instead  of 
vertical  and  another  saw  introduced,  though  these  do  not  both 
operate  at  the  same  time.  This  particular  outfit  is  driven  by 
an  old  7  X  10-in.  engine,  connected  up  as  shown. 


Description  of  Test 

MAXIMUM 
LOAD 

MAXIMUM 
SETTLEMENT 

Total 
Pounds 

Pounds  per 
Square 
Foot 

Inches 

Per  Cent 

Rectangular  pillar  of  mine  rock.  .  .  . 
Timber  crib  filled  with  mine  rock  .  .  . 
Circular  pillar  of  mine  rock  
Pile  of  broken  sandstone,  small  sizes  . 
Pile  of  broken  sandstone,  large  and 
small  sizes           

489,150 
900,000 
361,600 
581,000 

417,000 
600,000 

800,000 
200,000 
300,000 

300,000 
300,000 
300,000 
300,000 
600,000 

42,000 
63,300 
85,000 
209,000* 

150,000* 
216,000* 

228,000* 
886,000 
1,330,000 

1,330,000 
1,330,000 
1,330,000 
1,330,000 

* 

5.26 
7.08 
4.51 
4.36 

4.61 
5.00 

4.69 
2.73 
3.66 

2.42 
5.33 
3.00 
3.35 
2.43 

29 
30 
31 
46 

41 
63 

45 
30 
35 

23 

51 
33 
32 

27 

Pile  of  river  sand 

Pile  of  small  broken  sandstone  and 
sand 

Wet  culm  in  cylinder  
Broken  sandstone  in  cylinder  
Broken  sandstone  and  sand  in  cylin- 
der   
Cinders  in  cylinder           

Wet  culm  in  cylinder 

River  sand  in  cylinder  

Pillar  of  mine  gob 

*  Pressure  under  20X20  in.  bearing  plate. 

The  posts  are  first  sawed  on  the  slabber,  and  then  squared 
on  the  ends  by  running  through  the  machine,  say  one  hundred 
of  them,  and  the  saw  is  then  raised  up  so  as  to  cut  just  2  in. 
deep.  The  horizontal  saw  remains  stationary,  the  stick  is 
shoved  through  and  the  cut  made  on  top ;  the  carriage  is  then 
shoved  on  the  horizontal  saw  and  a  slice  taken  out  of  the  bot- 
tom; it  is  then  pulled  back  and  rolled  over  one  quarter  and 
the  operation  repeated;  it  requires  four  cuts  to  finish  the  end. 


378 


COAL  MINING  COSTS 


TIMBERING  COSTS  379 

The  complete  operation  averages  3y2  min.  to  frame  both  ends 
of  an  8  X  8-in.  post.  The  slabbing  saw  takes  logs  up  to  8  ft. 
in  length  and  the  company  made  a  set  of  dogs  and  some  per- 
forated plates  to  hold  pins,  and  bought  an  inserted  tooth  saw 
to  replace  the  thin  saw  which  came  with  it. 

The  machine  can  also  be  used  for  sawing  straight  lumber 
of  any  dimensions  up  to  8  ft.  in  length,  this  particular  instal- 
lation reducing  lumber  costs  at  the  mine  about  one-half.  A 
small  wedge  saw  was  added  for  cutting  scraps  into  wedges 
which  were  made  at  a  cost  of  %c.  each  as  compared  with  a 
cost  of  6c.  when  made  by  hand.  Some  of  the  other  advantages 
of  the  machine  are  that  the  slabs  made  can  be  generally  used 
for  lagging  in  the  less  critical  places  and  the  machine  fram- 
ing has  been  found  more  accurate  and  giving  better  fits  in  the 
mine. 

Timber  preservatives. — In  1906  the  United  States  Forest 
Service,  in  cooperation  with  the  Philadelphia  &  Reading  Coal 
&  Iron  Co.  carried  on  a  series  of  experiments  to  determine 
the  best  method  of  prolonging  the  life  of  mine  timber.  It  was 
found  as  a  result  of  these  studies  that  45  per  cent  of  the  mine 
timber  is  destroyed  by  decay,  while  breakage,  wear  and  insects 
accounted  for  the  balance.  Germs  or  spores  which  produce 
decay  may  gain  access  to  the  timber  at  any  time  before  or  after 
it  is  cut,  though  for  the  most  part  the  disease  is  contracted 
in  the  mines  from  decaying  timber  near  by.  In  untreated 
timber,  rough  surfaces  of  bark  and  wood  furnish  a  foothold 
for  the  spores,  which  subsequently  germinate  and  attack  the 
wood  tissues.  Spores  may  also  enter  timber  only  superficially 
treated  through  checks,  cracks,  or  nail  wounds. 

For  a  fungus  to  exist  it  must  have  a  definite  amount  of  air 
and  water,  food,  and  heat.  If  mining  conditions  were  such 
that  the  timber  would  be  kept  always  wet  or  always  dry,  it 
would  never  decay.  It  is  the  alternating  wet  and  dry  condi- 
tions or  continuous  dampness  which  produce  rot.  Tentilation 
is  a  very  large  factor  in  the  life  of  mine  timber.  Poorly  ven- 
tilated gangways  and  air  passages,  with  a  fair  degree  of  mois- 
ture and  a  fairly  high  temperature,  are  favorable  to  fungous 
growth,  and  hence  to  rapid  decay.  It  is  probably  impossible 
to  exterminate  disease,  sufficiently  to  wholly  prevent  decay  in 
mine  timber.  Right  preservative  treatment,  together  with  care- 


380  COAL  MINING  COSTS 

ful  handling  of  the  timber  will,  however,  reduce  both  to  a 
minimum. 

The  important  part  which  insects  play  in  the  destruction  of 
mine  timber  is  seldom  realized.  They  are  for  the  most  part 
brought  into  the  mines  with  the  timber.  Regular  and  thorough 
inspection  and  the  rigid  condemnation  of  insect-infested  timber 
would  therefore  greatly  reduce  the  loss  from  this  source. 

Insects  bore  into  the  sound  wood  and  greatly  weaken  it 
and,  moreover,  leave  holes  which  encourage  the  entrance  of 
wood-destroying  fungi.  A  good  preservative  treatment  will 
protect  the  timber  from  insect  attack,  as  well  as  prevent  decay. 
If  the  bark  is  removed  from  the  timber  soon  after  it  is  cut, 
it  will  not  be  attacked  by  wood-destroying  insects  until  the 
wood  becomes  old  arid  dry,  after  which  it  may  be  attacked  by 
"powder  post"  and  other  borers. 

Sets  of  round  gangway  timber  averaging  13  in.  in  diameter 
were  chosen  as  the  basis  for  the  experimental  treating  work. 
These  sets  in  the  anthracite  regions  consist  of  two  legs,  usually 
9  to  10  ft.  long  and  a  collar  6  to  7  ft.  long,  and  they  are  usually 
placed  on  5-ft.  centers.  The  sets  contain  26  cu.  ft.  of  timber 
and  the  average  life  in  the  anthracite  mines  is  two  years. 

Experiments  have  shown  that  peeled  timber  is  superior 
in  durability  to  unpeeled  timber.  The  space  between  the  bark 
and  the  wood  especially  favors  the  development  of  wood- 
destroying  fungi,  and  is  a  breeding  place  for  many  forms  of 
insect  life.  When,  after  placement  in  the  mines,  the  bark 
begins  to  flake  off,  the  timber  has  already  begun  to  decay. 
The  cost  of  peeling  timber  before  it  goes  into  the  mine  ranges 
from  20c.  to  50c.  per  ton  of  wood  (figures  as  of  1905),  accord- 
ing to  local  conditions  and  the  kind  of  timber. 

Seasoning  or  drying  gives  mining  timber  greater  strength 
and  durability.  A  stick  of  wet  timber  has  only  one-half  the 
strength  it  has  when  thoroughly  dry.  Though  it  is  not  prac- 
ticable for  mining  companies  to  hold  their  timber  until  it  is 
absolutely  air  dry,  peeled  timber  will  dry  out  sufficiently  in 
a  few  months  to  gain  in  both  strength  and  durability.  From 
two  to  four  months  is  necessary  for  proper  seasoning. 

To  determine  the  possible  loss  in  weight  in  round  timber, 
due  to  peeling  and  seasoning,  a  test  was  conducted  at  one  of 
the  collieries  of  the  company.  Representative  sticks  of  South- 


TIMBERING  COSTS 


381 


ern  loblolly  pine,  averaging  11  to  13  in.  in  diameter  and  from 
9  to  10  ft.  in  length  were  chosen.  This  timber  was  weighed 
immediately  before  and  after  peeling,  to  determine  the  weight 
of  the  bark.  It  was  then  weighed  every  two  weeks  until, 
seasoned,  to  learn  the  weight  of  the  water  evaporated.  The 
time  of  the  year  greatly  favored  rapidly  seasoning.  The  short 
lengths  into  which  the  timber  was  sawed  gave  a  large  drying 
surface  in  proportion  to  volume  and  longer  sticks  would  season 
more  slowly.  The  accompanying  diagram,  Fig.  2,  shows  the 


I 

i, 


I*       Z&        42        56        70       84-       98 
Number  of  Days  Seasoned 

FIG.  2. — Percentage  of  loss  of  green  weight  in  seasoning. 

average  percentage  of  loss  from  the  green  weight  in  seasoning, 
a  synopsis  of  the  results  of  the  tests  being  as  follows : 

PEELING  AND  SEASONING  TEST,  SOUTHERN  LOBLOLLY  PINE  ROUND  TIMBER? 
APRIL  17  TO  JULY  24,  1906 

Per  Cent 

Total  loss  of  green  weight  by  peeling 8.1 

Total  loss  of  green  weight  by  seasoning 35 . 1 

Peeling  and  seasoning 43.2 

RATE  OF  SEASONING 


Number  of  Days 

Percentage  of  Green 

Number  of  Days 

Percentage  of  Green 

Seasoned 

Weight  Lost 

Seasoned 

Weight  Lost 

14 

16.2 

70 

31.4 

28 

20.5 

84 

33.7 

42 

26.2 

98 

35.1 

56 

30.3 

If  a  mining  company  handles  its  own  timber  from  the  woods 
to  the  mines,  the  saving  in  freight  made  possible  by  peeling 


382 


COAL  MINING  COSTS 


and  seasoning  can  readily  be  estimated.  Labor  is  the  principal 
factor  in  the  cost  of  peeling,  while  the  cost  of  seasoning  must 
be  represented  by  the  loss  of  interest  on  the  capital  invested 
in  the  timber  during  the  seasoning  period.  However,  these 
additional  items  of  expense  are  more  than  offset  by  a  maxi- 
mum reduction  in  freight  of  from  30  to  40  per  cent  and  by 
the  far  better  condition  of  the  timber  with  regard  to  both  its 
life  at  the  mines  and  the  readiness  with  which  it  will  take 
preservative  treatment.  The  peeling  of  timber  at  the  mines  has 
been  unsatisfactory  and  expensive,  because  of  the  limited 
amount  of  yard  room  and  the  accumulation  of  bark.  The 
following  considerations  favor  peeling  in  the  woods:  (1)  The 
saving  in  the  cost  of  freight  due  to  peeling  and  seasoning;  (2) 
the  saving  of  yard  room  at  the  mines;  and  (3)  the  prevention 
of  fungous  disease  and  insect  attack  by  early  peeling. 

Peeling  and  seasoning  mine  timber  unquestionably  increase 
its  durability.  However,  in  order  to  prolong  its  life  to  the 
fullest  extent,  a  preservative  treatment  is  necessary.  The  in- 
creased life  necessary  to  justify  the  cost  of  applying  a  preserva- 
tive by  the  several  methods  in  vogue  is  indicated  by  the 
accompanying  diagram,  Fig.  3. 


Creosote 

^j 

fl 

BRUSH  • 

if 

Soft  Sot  uf  ion 

7 

TANK       • 

a 

41 

?7 

CYLINDER 

Creosote 

; 

SJ 

FIG.  3. — Increased  life  necessary  to  pay  cost  of  preservation  treatment. 

Impregnated  wood  resists  decay  because  the  preservative 
is  antiseptic  and  excludes  the  moisture  necessary  for  fungous 
growth.  Timber  used  in  mines  was  treated  with  a  variety  of 
preservatives  under  several  methods  of  application.  Both  green 
and  seasoned  timbers  were  treated  to  determine  both  the  rela- 
tive value  of  the  treatments  and  the  best  method  of  handling 
preparatory  to  treatment.  If  treated  at  all  the  timber  musl 
be  peeled.  The  accompanying  table  shows  the  method  of  treat- 


TIMBERING  COSTS 


383 


merit,  the  preservative  applied,  the  cost  of  same  and  the  cost 
of  the  treatment  for  an  average  gangway  set  and  per  cubic 
foot: 


COST  OF 

TREATMENT 

Absorp- 

Cost of 

tion 

Method  of 

Preservative 

Pre- 

Per Set 

per 

Treatment 

Applied 

servative 

of 

Ppr 

Cubic 

Gangway 
Timber 

JT  ci 

Cubic 

Foot, 

Foot 

(25.8 

Per  gal. 

Cu.  Ft.) 

Pounds 

f  Creosote  (dead  oil  of 

Brusn  

<       coal  tar)  

$0.09 

$0.40 

$0  .  015 

1  Carbolineum  

.70 

1.15 

.045 

fSalt  solution,  magne- 

sium,    chloride    15 

Open  tank  with- 

per cent  

.01 

.50 

.020 

out  pressure  . 

Zinc  chloride  solution, 

6  per  cent 

.04 

.90 

.035 

1  Creosote  

.09 

2.85 

.110 

10 

Cylinder      with 
pressure  

(Zinc  chloride  solution, 
6  per  cent  
Creosote  

.04 
.09 

1.90 
3.85 

.075 
.15 

10 

Brush  treatments  with  both  creosote  and  carbolineum  were 
applied  in  two  coats  to  the  Pennsylvania  and  Southern  pines. 
A  large  flat  brush  and  kettle  of  the  hot  preservative  are  all 
that  is  required  for  this  treatment.  A  very  small  amount  of 
the  preserving  fluid  suffices,  but  the  cost  of  application  in  pro- 
portion to  the  results  obtained  is  considerable.  For  small 
individual  operators  who  cannot  afford  the  cost  of  a  large 
plant,  brush  treatments  are  feasible  and  economical. 

The  disadvantages  of  brush  treatments  are : 

(1)  The  difficulty  of  completely  covering  the  timber  and 
filling  all  checks  and  cracks. 

(2)  The  very  slight  penetration  secured.     The  subsequent 
checking  or  opening  of  the  timber  may  often  allow  disease  to 
pass  through  the  shallow  exterior  band  into  the  untreated 
interior  wood. 


384 


COAL  MINING  COSTS 


Pitch  pine  and  loblolly  pine  have  been  most  successfully 
treated  with  both  creosote  (dead  oil  of  coal  tar)  and  a  6-per 
cent  solution  of  zinc  chloride  by  the  open-tank  process. 

The  open-tank  treatment  as  given  in  this  experiment  was 
briefly  as  follows:  Green,  partially  seasoned,  and  thoroughly 
seasoned  timber  was  lowered  into  the  tank  and  immersed  in 
creosote,  or  in  a  zinc  chloride  or  salt  solution,  at  a  temperature 
of  from  90  deg.  to  120  deg.  F.  The  temperature  of  the  creosote 
was  raised  by  the  coils  to  from  212  deg.  to  220  deg.  F.,  and 
that  of  the  zinc  chloride  or  the  salt  solution  to  about  212  deg. 
F.  In  no  case,  however,  was  the  temperature  allowed  to  go 
above  240  deg.  F.  for  fear  of  injuring  the  fiber  of  the  timber 
and  so  decreasing  its  strength.  When  this  hot  bath  was  over 
•the  steam  was  turned  off,  and  the  timber  was  allowed  to  stand 
until  the  liquid  cooled  to  a  temperature  of  from  170  deg.  to 
^100  deg.  F.  The  periods  of  heat  and  of  cooling  were  varied 
for  each  kind  of  timber  and  for  each  stage  of  its  seasoning. 
The  time  required  for  the  cooling  operation,  which  depended 
largely  upon  the  temperature  of  the  atmosphere,  was  usually 
from  3  to  12  hr.  For  the  whole  treatment  the  time  varied  from 
6  to  20  hr. 

Loblolly  and  pitch  pine  can  be  successfully  and  economically 
treated  by  simple  immersion  in  successive  hot  and  cold  baths 
in  an  open  tank,  at  a  cost  of  about  lie.  per  cubic  foot.  Green 
timber  is  treated  with  far  more  difficulty  than  seasoned  timber. 


AVERAGE  AND  REPRESENTATIVE  TREATMENTS  OF  LOBLOLLY  AND  PITCH  PINE 
BY  THE  OPEN-TANK  PROCESS 

CREOSOTE 
AVERAGE  ABSORPTION  AND  PENETRATION,  LOBLOLLY  PINE    (PINUS  T/FDAI) 


Absorption 

Depth  of 

Condition  of  Timber 

per  Cubic  Foot, 

Penetration, 

Pounds 

Inches 

Green  

5-  7 

T-li 

Seasoned  (1  to  2  months)     .               ... 

12-15 

2  -4 

Seasoned  (3  to  4  months)  

20-25 

5-complete 

TIMBERING  COSTS 


385 


REPRESENTATIVE  INDIVIDUAL  RUNS,   SEASONED  LOBLOLLY  PINE   (NEARLY 

ALL  SAP  WOOD) 


Time 
Seasoned, 

Total 
Length  of 
Treat- 
ment, 

Duration 
of 
Hot  Bath, 

TEMPERATURE 

Absorption 
per 
Cubic 
Foot, 

Penetra- 
tion, 

Average, 

Maxi- 

Months 

Hours 

Hours 

°F. 

°F. 

Pounds 

Inches 

3 

24 

7 

230 

240 

22.0 

4-5 

3 

24 

4| 

225 

235 

21.5 

4-5 

3 

6* 

if 

178 

220 

10.7 

2-3 

3 

6 

2 

173 

210 

10.7 

2-3 

3 

H 

11 

174 

198 

10.2 

2-3 

REPRESENTATIVE  INDIVIDUAL  RUNS,  SEASONED  PITCH  PINE   (HEART  WOOD 

AND  SAP  WOOD) 


Time 
Seasoned, 

Total 
Length 
of 
Treat- 
ment, 

Duration 
of 
Hot 
Bath, 

TEMPERATURE 

. 

Absorp- 
tion, 
Pounds 
per 
Cubic 

Pene- 
tration, 

Width 
of 
Sap 
Wood, 

Aver- 
age, 
o  jp 

Maxi- 
mum, 

O    "IT* 

" 

Font 

Months 

Hours 

Hours 

-1?  UU  L 

Inches 

Inches 

4 

22 

71 

215 

240 

6| 

3 
4 

1 

4 

22 

7£ 

218 

240 

12 

If 

11 

4 

22 

7| 

209 

232 

2U 

3 

2| 

i 

i 

SOLUTION  OF  ZINC  CHLORIDE  (6-8  PER  CENT) 
THOROUGHLY  SEASONED  LOBLOLLY  PINE  (NEARL\  ALL  SAP  WOOD) 


Total 

Length  of 

TEMPERATURE 

Length  of 
Treatment, 

Period  in  Hot 
Solution, 

Absorption, 

Pene- 
tration, 

Average, 

Maximum, 

Pounds  per 

Hours 

Hours 

°F. 

°F. 

Cubic  Foot 

Inches 

20 

6 

200 

210 

20 

3-5 

20 

6 

200 

210 

35 

4-6 

386  COAL  MINING  COSTS 

The  difference  in  weight  of  green  timber  before  and  after 
treatment  is  by  no  means  indicative  of  the  amount  of  the 
preservative  absorbed.  The  simple  application  of  the  hot 
liquid  to  green  timber  slightly  reduces  its  weight  and  yields 
no  penetration.  The  same  application  to  seasoned  timber 
slightly  increases  its  weight  and  gives  a  slight  penetration. 
Green  timber  after  treatment  may  show  a  penetration  of  1  inch 
without  an  increase  in  weight. 

Heart  wood  of  both  loblolly  pine  and  pitch  pipe  is  pene- 
trated with  far  more  difficulty  than  is  the  sap  wood  of  the 
same  species.  This  is  especially  the  case  with  pitch  pine  which 
clearly  shows  after  treatment  a  distinct  division  between  the 
treated  sap  wood  and  the  untreated  heart  wood. 

Experiments  indicate  that  for  pine  timbers  of  the  same 
degree  of  dryness,  or  containing  equal  proportions  of  heart 
wood  and  sap  wood,  impregnation  can  be  regulated  by  increas- 
ing or  decreasing  the  duration  of  the  cooling  bath. 

During  the  year  1906-7  gangway  timber  of  various  species, 
peeled  and  unpeeled  green  and  seasoned,  and  treated  and 
untreated  was  placed  in  gangways  in  the  collieries  of  the  com- 
pany. Each  and  every  kind  and  condition  of  gangway  timber 
has  been  compared  with  the  timber  in  most  general  use  in  the 
southern  anthracite  region,  namely,  green,  unpeeled  loblolly, 
and  pitch  pines.  The  object  of  this  comparison  is  to  prove 
exactly  to  what  extent  the  experimental  timber  is  superior  to 
that  at  present  used.  In  the  course  of  the  experimental  work 
the  following  comparisons  have  been  made: 


Species  Compared 


Treatments  Compared 


Loblolly  pine  (Pinus  Tceda)  .    \   Creosote 

Brush 


Pitch  pine  (Pinus  rigida) 
Longleaf  pine  (Pinus  palustris) 
Chestnut  (Castanea  dentata) 
Red  oak  (Quercus  rubra) 


Carbolineum 

{Creosote 
Solution  of  zinc  chloride 
Solution  of  sodium  chloride  and 
magnesium  chloride 

-,  ,.  ,    Creosote 

Cylinder  S    „  i     •         r    •         i  i     •  i 
Solution  of  zinc  chloride 


The  history  of  each  set  of  gangway  timber  and  each  part 
of  each  set  has  been  recorded  in  writing  and  on  maps.  These 
records  include:  (1)  The  date  of  setting;  (2)  the  colliery; 


TIMBERING  COSTS 
SUMMARY  OF  EXPERIMENTAL  SETS  OF  TIMBERS 


.  387 


COLLIERY 

Silver  Creek 

Eagle  Hill 

Wadesville 

Total 

Untreated  : 

11  loblolly 

26  loblolly 

37 

44  loblolly 

16  loblolly 

98 

Green  unpeeled  

^  16  longleaf 
'  112  loblolly 
31  pitch  pine 

8  longleaf 
)36  loblolly 

f  26  pitch  pine 

221 

8  black  oak 

7  pitch  pino 

Total  untreated  

356 

Brush  treatment: 
Green  — 

14  loblolly 

9  loblolly 

27 

f  18  loblolly 

>  9  loblolly 

38 

Seasoned  — 
Carbolineum  

\  5  chestnut 

7  loblolly 
9  loblolly 

9  loblolly 
28  loblolly 

6  pitch  pine 
5  loblolly 

22 
42 

Total  brush  treatment.  . 

129 

Tank  treatment: 
Green  —  creosote  

(104  loblolly 
7  chestnut 

I  6  loblolly 

124 

Seasoned  — 

5  pitch  pine 
2  black  oak 

f  20  loblolly 

31 

Salt  
Zinc  chloride  

\  11  pitch  pine 
f  6  loblolly 

f  17  pitch  pine 
\  14  loblolly 

}" 

11 

\  5  longleaf 

Total  tank  treatment.  .  . 

97 

Cylinder  treatment,  seasoned: 
Creosote  

23  loblolly 

23 

Zinc  chloride  

50  loblolly 

50 

Total  cylinder  treatment.  . 

73 

Grand  total   . 

755 

(3)  the  gangway;  (4)  the  position  in  the  gangway  relative  to 
the  nearest  chute  and  adjacent  set  of  timber. 

The  accompanying  table  gives  a  summary  of  the  experi- 


388 


COAL  MINING  COSTS 


mental  sets  of  timber  placed  in  the  mines  of  the  Philadelphia 
&  Beading  Coal  and  Iron  Co.  in  1906. 

PRESERVATIVE  TREATMENT  APPLIED 


Method  of  Application 

Preservative 
Used 

Approximate 
Cost  of 
Preservative 

Approximate 
Cost  per  Set 
of  Gangway 
Timber  of 
26  Cu.  Ft. 

Cost 
per 
Cubic 
Foot 

The  preservative  heated  to 
180°  F.  and  applied  in  two 
coats  with  a  brush 

Creosote    (dead   oil 
of  coal  tar)  
Avernarius     carbo- 
lineum        

$0.09    per  gal. 
0.70    per  gal. 

$0.40 
1.15 

$0.01* 
0.04* 

Immersion  in  an  open  tank 
without  pressure  —  succes- 
sive baths  of  hot  and  cold 
fluid.  Plant  of  simple  con- 
struction 

Solution  of  common 
salt  (15  per  cent) 
Solution      of      zinc 
chloride     (6    per 

$0.009  per  Ib. 

0.04}  perlb. 
0  .  09    per  gal. 

$0.50 

0.90 

2.85 

$0.02 

0.03* 
0.11 

Creosote    (dead   oil 
of  coal  tar)  

In  a  closed  cylinder  under  vac- 
uum and  pressure.  Plant  of 
complex  construction 

Solution      of      zinc 
chloride      (6    per 
cent)  
Creosote    (dead   oil 
of  coal  tar)  

$0.04}  perlb. 
0.09    per  gal. 

$1,90 
3.85 

$0.07 
0.15 

Steel  timbering. — The  following  are  the  comparative  costs 
of  steel  and  frame  timbering  as  found  under  actual  working- 
conditions  : 

In  1908  at  their  Maxwell  colliery  the  Lehigh  &  Wilkes- 
Barre  Coal  Co.  timbered  a  double-track  gangway  with  20  in. 
65-lb.  I-beam  collars  17  ft.  long  between  supports,  and  8  in. 
H-beam  legs  10  ft.  6  in.  high  in  the  clear,  weighing,  with  base 
plates,  1720  Ib.  per  set.  These  took  the  place  of  wooden  sets 
made  of  24  in.  round  yellow  pine  timbers,  the  cost  of  which 
erected  was  $15  per  set,  weight  5040  Ib.  and  the  life  of  which 
was  two  and  one-half  years.  In  view  of  their  probable  dura- 
bility, the  steel  sets  were  erected  on  concrete  bases  and  this 
added  to  the  cost  of  installation,  which  reached  a  total  of 
$40  per  set. 

Capitalized  at  6  per  cent  interest,  the  value  of  the  steel 
sets  at  the  end  of  15  yr.  will  be  $95.86  each,  while  the  capital- 
ized value  of  the  six  wooden  sets  needed  in  that  time  will  be 


TIMBERING  COSTS  389 

$153.56.  At  the  end  of  the  15  yr.,  the  steel  will  have  a  scrap 
value  per  set  of  $12.03,  while  the  wood  will  be  worth  nothing, 
a  saving  by  the  use  of  steel  of  $69.73  per  set  or  $4.65  per  year. 
The  pump  house  at  the  Dodson  colliery,  of  the  Plymouth 
Coal  Co.,  is  100  ft.  long,  8  ft,  high  in  the  clear  and  18  to 
22  ft.  wide.  The  roof  is  exceedingly  tender  and  has  caused 
all  kinds  of  trouble  in  the  pump  house,  especially  in  connec- 
tion with  the  pipes.  Before  retimbering  with  steel,  18  to  22  in. 
round  timbers  were  used,  on  2-ft.  centers,  practically  skin  to 
skin.  It  is  estimated  that  the  pump  room  was  retimbered  once 
a  year.  Beginning  with  April,  1910,  the  70  sets  of  wood  tim- 
bers were  replaced  by  48  sets  of  steel.  The  last  steel  set  was 
placed  December  15,  1910.  According  to  figures  furnished  by 
John  C.  Haddock,  president  of  the  company,  the  relative  costs 
were: 

1.  Wood — 70  sets;  weight  per  set  4150  Ib. ;   cost  per  set 
f.o.b.  cars  at  mine,  $12;  cost,  erected  in  place,  $34.50;  total  cost, 
$2415  ;  life,  one  year. 

2.  Steel — 48  sets ;  weight  per  set  1483  Ib. ;  cost  per  set  f.o.b. 
cars  at  mine,  $31.47;  cost  erected  in  place,  including  concrete 
footings,  $61.47  per  set;  total  cost  $2889.09,  or  not  quite  20  per 
cent  more  than  wood. 

Based  on  its  life,  the  cost  per  month  of  a  wood  set  without 
interest  was  $201.25.  The  cost  of  the  steel  sets  at  the  end  of  16 
months  without  interest  was  $180.57  per  month.  At  the  end 
of  that  time  they  had  shown  no  signs  of  failure. 

At  the  No.  8  mine  of  the  West  Kentucky  Coal  Co.,  steel 
timbers  are  used  in  a  slope,  both  for  the  main  entry  and  air 
course.  The  sets  are  composed  of  a  10  in.  25-lb.  I-beam  collar 
and  4  in.  H-beam  legs.  They  are  spaced  3  ft.  centers  and 
lagged  with  oak  plank  3  in.  thick  on  top,  and  2  in.  thick  on  the 
sides.  Between  the  sets,  concrete  is  placed  up  to  a  height  of 
4  ft.  This  makes  a  solid  reinforced  concrete  slope  from  the 
entrance  to  the  point  where  the  ribs  are  hard  and  top  good. 
According  to  figures  furnished  by  W.  H.  Cunningham,  general 
manager  of  the  company,  the  comparative  costs  of  wood  and 
steel  for  his  mine  were  (about  1912)  : 

Wood— Yellow  pine  creosoted ;  size  12  X  12  in.,  264  ft.  b.m. ; 
cost  at  Sturgis,  $10.56  per  set;  cost  in  place,  $15.70;  weight 
1575  Ib. 


390  COAL  MINING  COSTS 

Wood— Native  white  oak;  size  12X12  in.,  264  ft.  b.m.; 
cost  at  Sturgis,  $7.92;  cost  in  place,  $13.06  per  set;  weight 
1340  Ib. 

Steel — Cost  of  steel  at  Sturgis,  $9.75  per  set;  cost  of  plac- 
ing $1;  cost  of  concrete  per  panel  $5.16;  total  cost  in  place 
per  set,  steel  alone  $10.75,  steel  concreted  $15.91 ;  weight  of 
steel  sets  425  Ib. 

Saving  in  the  use  of  steel  without  concrete,  over  native 
white  oak,  $2.31  per  set,  over  yellow  pine  $4.95.  Excess  cost 
of  steel  with  concrete,  over  white  oak  $2.85  per  set,  over  yellow 
pine  21c.  This  favorable  comparison  is  due  to  the  high  unit 
cost  of  the  wood  and  to  the  elimination  of  waste.  The  safe 
uniformly  distributed  load  on  the  wood  collar  is  1200  Ib.,  on 
the  steel  collar  26,000  Ib.  The  safe  compressive  strength  of  the 
steel  leg  is  43,200  Ib.,  while  that  of  the  wooden  leg  is  105,100 
Ib. ;  in  the  one  case  more  than  ample,  in  the  other  case  out  of 
all  proportion. 

In  some  cases  it  has  even  been  found,  where  transportation 
costs  are  not  excessive,  that  the  first  cost  of  the  steel  timber- 
ing erected  in  place  is  equal  to  or  but  slightly  more  than  the 
cost  of  a  wooden  installation  of  similar  strength.  It  has  also 
been  found  that  it  is  possible  to  fit  the  steel  exactly  to  the 
structural  necessities,  whereas  it  is  impracticable  in  all  cases 
to  do  so  with  wood.  This  circumstance  eliminates  an  economic 
waste  due  to  the  use,  for  practical  considerations,  of  larger 
sized  sticks  of  wood  than  are  really  necessary. 

The  accompanying  table  was  prepared  by  a  mining  company 
in  western  Illinois  and  compares  the  relative  cost  of  steel  beam 
collars,  square-sawed  white  oak  beams  and  square-sawed  yel- 
low-pine beams  delivered  underground.  It  also  estimates  the 
comparative  maintenance  costs  for  a  period  of  20  yr.,  based 
on  the  most  favorable  and  the  least  favorable  probable  con- 
ditions. The  safe  working  loads  are  based  on  the  1915  values 
for  the  materials  under  bending  stress. 

The  smaller  table  gives  the  first  cost  of  such  beams  and 
their  cost  at  the  end  of  the  20-yr.  period  under  the  conditions 
assumed.  It,  therefore,  represents  what  might  be  called  a 
reasonable  expectation  of  relative  service  in  a  particular  dis- 
trict where  the  conditions  as  to  transportation,  cost  of  steel 
and  the  relative  availability  of  wood  are  normal. 


TIMBERING  COSTS 


391 


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COAL  MINING  COSTS 


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TIMBERING  COSTS  393 

Examination  of  this  table  will  indicate  that  in  the  case 
of  the  16-ft.  spans  there  is  but  little  difference  in  the  first  cost 
between  steel  and  white  oak,  and  that  at  the  end  of  20  yr.  under 
any  circumstance  steel  is  the  most  economical. 

This  economical  advantage  of  steel  as  compared  with  wood 
would  be  somewhat  enhanced  if  consideration  were  had  to  the 
matter  of  interest  on  the  investment,  which  however,  the  min- 
ing company  which  prepared  the  table  did  not  consider.  It 
can  also  be  said  that  repainting  of  steel  once  in  4  yr.  is  about 
the  limit  of  probability.  There  are  installations  that  have 
been  in  place  two  and  three  times  as  long  as  that  without 
repainting,  and  it  must  not  be  forgotten  that,  except  in  special 
locations,  conditions  underground  as  regards  paint  are  much 
better  than  they  are  above  ground. 

Steel  beams  for  roof  supports  may  be  used  just  as  they 
come,  cut  to  length,  from  the  rolling  mill;  all  the  other  types 
of  mine  timbering  need  to  be  fabricated,  that  is,  framed  and 
fitted,  before  they  are  ready  for  use  in  the  mines.  A  most 
important  consideration  is  to  reduce  the  cost  of  this  fabrica- 
tion to  the  lowest  possible  amount  by  simplification  of  the 
details.  Blacksmith  work  always  should  be  avoided ;  and  that 
form  of  framing  is  the  cheapest  in  which  the  fitting  shop  work 
is  the  simplest.  Intelligent  skill  in  structural  design  here  means 
great  saving. 

Eight  different  types  of  framing  for  steel  gangway  sup- 
ports have  come  into  serious  consideration.  With  plain  ma- 
terial at  1.25c.  per  pound  f .o.b.  cars  Pittsburgh,  and  the 
usual  extras  for  workmanship,  the  comparative  costs  of  these 
styles  are  shown  in  the  accompanying  tables  and  from  which 
can  be  seen  at  a  glance  how  great  a  part  attention  to  details 
may  play  in  the  economic  use  of  materials  and  the  avoidance 
of  unnecessary  work  in  fabrication.  The  eight  sets  of  each 
table  are  all  of  equivalent  theoretical  strength. 

The  figures  in  these  two  tables  do  not  include  painting. 
The  cost  for  this  varies  considerably  with  the  kind  of  paint 
used  but  may  be  estimated  at  $2  per  ton  (1912). 

Concrete  timber. — Many  modern  mines  are  using  steel  for 
supporting  gangways,  and  satisfaction  has  attended  these 
examples.  Even  where  loads  are  not  great  and  the  fire  hazard 
low,  the  use  of  metal  in  place  of  wood  has  been  found  a  profit- 


394 


COAL  MINING  COSTS 


STEEL  GANGWAY  SUPPORTS  FOR  A  DOUBLE  TRACK  GANGWAY 

Collar  17  ft.  long  between  legs;  legs  10  ft.  6  in.  high  in  the  clear;  equivalent 
in  strength  to  24  in.  round  longleaf  yellow  pine  timbers. 


WEIGHT  PER  SET 

COST  PER  SET 

Style 

Size  of  Collar 

Size  of  Pegs 

Without 

With 

Without 
Base, 

With 
Base 

Base, 

Base, 

Dollars 

Plates, 

Lb. 

Lb. 

per 

Dollars 

Set 

per  Set 

B 

20"  65-lb.  beam 

2-7"  C.  —  14.75  Ib. 

1930 

2030 

36.82 

39.48 

D 

20"  65-lb.  beam 

2-7"  C.—  14.75  Ib. 

1930 

2100 

36.82 

43.95 

C 

20"  65-lb.  beam 

1-8"  H.—  34.6    Ib. 

1710 

1800 

25.39 

27.80 

A 

20"  65-lb.  beam 

2-7"  C.—  14.75  Ib. 

1930 

2500 

36.82 

58.50 

F 

20"  65-lb.  beam 

1-8"  H.—  34.6    Ib. 

1690 

1730 

25.81 

26.88 

.  I 

20"  65-lb.  beam 

1-8"  H.—  34.6    Ib. 

1730 

1780 

25.22 

27.12 

E 

20"  65-lb.  beam 

2-7"  C.—  14.75  Ib. 

1670 

1770 

25.94 

28.60 

G 

20"  65-lb.  beam 

1-8"  H.—  34.6    Ib. 

1690 

1730 

25.81 

26.88 

STEEL  GANGWAY  SUPPORTS  FOR  A  SINGLE  TRACK  GANGWAY 

Collar  10  ft.  long  between  legs;  legs  8  ft.  high  in  the  clear;  equivalent  in 
strength  to  15  in.  round  longleaf  yellow  pine  timbers. 


WEIGHT  PER  SET 

COST  PER  SET 

Style 

Size  of  Collar 

Size  of  Legs 

Without 

With 

Without 
Base 

With 
Base 

Base, 

Base, 

Dollars, 

Plates, 

Lb. 

Lb. 

per 

Dollars 

Set 

per  Set 

A 

10"  25-lb.  beam 

2-6"  C.—  10.5  Ib. 

765 

945 

14.81 

23.17 

D 

10"  25-lb.  beam 

2-6"  C.—  10.5  Ib. 

765 

800 

14.81 

15.85 

B 

10"  25-lb.  beam 

2-6"  C.—  10.5  Ib. 

765 

810 

14.81 

16.12 

E 

10"  25-lb.  beam 

2-6"  C.—  10.5  Ib. 

605 

660 

9.39 

10.87 

F 

10"  25-lb.  beam 

1-5"  H.—  18.7  Ib. 

569 

590 

8.84 

9.43 

G 

10"  25-lb.  beam 

1-5"  H.—  18.7  Ib. 

566 

587 

8.80 

9.39 

C 

10"  25-lb.  beam 

1-5"  H.—  18.7  Ib. 

565 

605 

7.91 

9.04 

I 

10"  25-lb.  beam 

1-5"  H.—  18.7  Ib. 

600 

360 

9.80 

11.11 

TIMBERING  COSTS  395 

able  investment  on  the  basis  of  the  ultimate  economy  in  the 
expenditure,  which  results  from  low  maintenance  charges  and 
infrequent  renewals. 

Inasmuch  as  steel,  however,  is  considerably  more  expensive 
than  reinforced  concrete,  and  as  the  latter  possesses  all  the 
advantages  of  the  former,  it  would  appear  that  reinforced 
concrete  would  be  well  adapted  to  mine  timbering.  Of  course 
it  might  be  contended  that  on  account  of  its  low  tensile  strength 
and  relatively  high  resistance  to  compression  and  crushing, 
that  concrete  would  not  be  as  suitable  as  steel,  which  has  a 
high  tensile  strength,  and  is  therefore  fitted  to  carry  large 
loads  over  wide  spans  with  a  minimum  ratio  of  dead  weight 
to  external  loading. 

In  most  mines,  however,  the  gangways  are  so  driven  that 
it  is  unnecessary  to  cover  wide  spans.  If  it  were  found  neces- 
sary to  meet  such  conditions,  a  more  complicated  system  might 
have  to  be  considered,  and  a  few  props  could  be  so  placed  as 
to  reduce  the  span  while  leaving  the  passages  as  unobstructed 
as  possible. 

Proper  design  for  reinforced  concrete  used  in  mine  timber- 
ing should  be  based  on  correct  principles  of  engineering.  Props 
could  be  planned  according  to  the  principles  of  columns  used 
in  the  erection  of  buildings,  as  shown  in  Fig.  4,  consisting  of 
rods  embedded  in  the  concrete  near  the  periphery.  These  are 
connected  by  means  of  ties,  or  flats,  hoop  iron  or  wire.  Thus 
the  radius  of  gyration  is  increased  and  the  rods  take  care  of 
the  tensile  stresses  which  occur  from  eccentric  loading  or  from 
deflection  of  columns.  The  horizontal  ties  prevent  the  buckling 
of  the  rods  and  increase  the  strength  of  the  concrete.  They 
form  a  hooped  column. 

The  size  and  dimensions  of  these  props  or  columns  must 
be  determined  by  the  nature  of  the  roofs  of  the  mine  they  have 
to  support  and  on  the  height  of  the  slope  or  entry  wherein 
they  will  be  used.  Above  ground  it  is  possible  to  calculate  the 
loading  and  the  stresses  involved  with  mathematical  correct- 
ness, but  underground  there  are  no  well  defined  rules  by  which 
for  example  the  strength  of  square  timber  sets  or  props  may 
be  calculated  other  than  those  given  "in  this  chapter. 

Regarding  the  construction  of  these  reinforced-concrete 
members,  inasmuch  as  suitable  sand  and  gravel  can  be  found 


396 


COAL  MINING  COSTS 


in  almost  any  locality,  it  would  only  be  necessary  to  ship  to 
the  mine  the  steel  and  cement.  The  props  and  collars  could 
be  made  in  the  quantities  desired  on  the  surface  in  close  prox- 
imity to  the  mine. 

The  method  of  procedure  followed  would  be  along  the  lines 
of  that  pursued  in  the  manufacture  of  concrete  telegraph  poles, 
and  the  form  work  would  be  similar  in  construction.  The 
lumber  used  for  forms  should  be  of  either  1-,  l1/^-  or  2-in. 
boards,  securely  braced  by  3  X  3-  or  4  X  4-in.  braces,  and  bolted 
in  order  to  prevent  bulging.  The  greatest  economy  is  secured 


Round  rods,  size  ; 
and  spacing  to  ' 
bedetermined  •$ 
by  design 

Flats,  size  and 
verticafspac- 
ingfobe 
determined 
by  design 


Concrete-- 


FIG.  4. — A  prop  designed  like  a  building  column. 


by  constructing  the  forms  so  that  they  can  be  used  over  and 
over  again.  Economy  may  also  be  gained  by  fastening  the  form 
work  together  with  the  minimum  amount  of  nailing. 

The  relative  proportions  of  strength  of  reinforced  concrete, 
steel  and  timber  are  based  on  the  ordinary  method  of  cal- 
culations, as  shown  in  Figs.  5  to  1,  and  the  use  of  their  equiva- 
lents produces  much  stronger  and  stiffer  mine  sets  than  the 
comparison  would  seem  to  indicate.  Stiffness  is  as  important 
as  strength,  and  the  spacing  should  be  such  as  to  compel  the 
different  sets  to  act  together  as  a  unit  under  any  sudden  stress 
or  shock.  Light  designs  with  a  close  spacing  will  therefore  be 
preferable  to  heavy  ones  with  wide  spacing,  the  roof  itself 
serving  as  a  beam  to  distribute  the  load  over  two  or  more 
sets,  whereas  on  wide  spacing  there  is  more  danger  of  the 


TIMBERING  COSTS  397 

roof  falling  in  between  sets.  The  closer  spacing  also  permits 
of  much  lighter  lagging. 

Conditions  in  the  mine  as  to  sizes  of  timbers  used,  the  near- 
ness to  a  supply  and  the  cost  of  lumber  vary  widely,  and  an 
infinite  number  of  comparisons  as  to  the  relative  costs  of  re- 
inforced concrete,  steel  and  wood  could  be  instituted,  results 
of  which  might  not  be  of  exact  value.  Under  market  condi- 
tions as  of  1914,  the  cost  of  steel  in  Pennsylvania  is  nearly 
three  times  the  cost  of  wood  used  in  the  square  timber  sets, 
and  the  cost  of  reinforced  concrete  may  be  taken  at  about  one 
and  one-third  times  the  cost  of  wood.  In  the  first  cost  of  the 
installation,  therefore  the  advantage  will  lie  with  wood. 

Where,  however,  a  gangway  has  to  be  maintained  over  a 
number  of  years  and  the  workings  are  in  any  way  permanent, 

i°-r- 

Reinforced- 
Concrete 

6*6"    .. 
jfeinforce'd 
Concrete 


Timber  Set  Reinforced  Concrete  Set 

FIG  5.  FIG  6. 

Comparative  size  of  timber,  steel  and  concrete  sets  of  equal  strength. 

consideration  should  be  given  to  the  capitalized  value  of  the 
material  as  compared  with  the  first  cost  of  the  installation ;  and 
reinforced  concrete  will  be  found  economical  in  most  cases  on 
the  basis  of  ultimate  cost  by  reason  of  its  long  life  and  endur- 
ance. 

Taking  as  a  basis  of  estimate  the  price  of  structural  steel 
at  $1.405  per  100  Ib.  f.o.b.  cars  Pittsburgh,  with  the  usual 
extras  for  workmanship,  etc.,  a  comparison  of  the  relative  costs 
of  the  form  of  gangway  supports  shown  in  Figs.  5  to  7,  is  as 
follows : 

The  cost  of  the  timber  sets,  as  shown  in  Fig.  5,  was  $7.50, 
erected.  The  cost  of  the  steel  sets,  as  shown  in  Fig.  7,  would 
be  $22,  erected.  If  painted,  the  cost  would  be  about  $2  per 
ton  additional.  The  cost  of  the  reinforced-concrete  set  would 
be  $10,  erected. 


398 


COAL  MINING  COSTS 


Reckoning  6  per  cent  compound  interest,  on  the  low  assump- 
tion of  15  yr.  life,  the  steel  set  at  the  end  of  this  period  will 
represent  an  investment  of  $52.72,  and  the  reinforced  concrete 
set  $23.97.  During  the  same  length  of  time,  based  on  past 
experience  and  on  the  present  cost  of  timber,  six  wooden  sets 
would  be  required  at  a  capitalized  value  of  about  $100. 

On  this  basis,  the  saving  on  the  steel  timbers  can  be  set 
down  as  $47.28  per  set,  which  means  a  saving  per  year  of  $3.15 ; 
on  the  reinforced  concrete,  the  saving  per  set  is  $76.03,  which 
amounts  to  a  yearly  saving  of  $5.07  over  lumber,  and  over 
steel  of  $1.92  per  year  per  set. 


0    "_ 


H- 


Ifocfe 

1 1-      \l' 


1 


Design  of  Collar 


Stirrups  of %"°/?ocfs 


Abob 


<p 

~±Flcrts,or 

\  u^ 

JL* 


Section  A~B 

Design  of  Prop 
FIG.  8. — Design  for  a  reinforced-concrete  timber  set. 

A  reinforced-concrete  design  is  shown  in  Fig.  8,  and  the 
above  cost  is  figured  as  follows : 

Steel  rods,  including  stirrups,  labor  and  material $2 . 25 

Aggregates,  sand,  cement  and  stone,  hauling,  mixing  concrete  and 

placing,  including  foreman's  wages 3 . 28 

*Forms,  carpenter  work,  labor  and  material  (assume  forms  used  three 

times,  equals  one-third  of  $6.25,  equals) 2.25 


$7.78 
Transporting  to  mine  and  erecting  (30%  of  $7.78) 2 . 22 


Total $10.00 

*  This  item  includes  removal  and  resetting  of  forms. 


TIMBERING  COSTS  399 

Regarding  the  handling,  the  following  is  a  comparison  of 
the  weights  per  set :  Timber  2200  lb.,  steel  650  lb.,  reinforced 
concrete  1800  lb.  From  these  figures  it  will  be  seen  that  the 
advantage  as  regards  weight  lies  with  steel,  but  this  handling 
would  only  figure  in  the  cost  of  transportation  from  the 
surface  in  proximity  to  the  drift,  slope,  or  shaft  mouth  to  the 
destination  in  the  mine,  and  would  not  in  any  way,  enter  into 
freight  charges,  the  transportation  costs  on  the  plain  concrete 
material  being  considerably  less  than  on  the  steel  set.  It  will 
also  be  seen  that  concrete  compares  favorably  in  weight  with 
lumber. 

At  shafts  No.  3  and  No.  4  of  the  copper  mine  of  the  Ahmeek 
Mining  Co.,  at  Ahmeek,  Mich.,  about  1912  the  use  of  concrete 
timbers  was  tried,  the  results  and  costs  of  which  are  given 
herewith : 

After  some  experimenting,  the  concrete  as  finally  used  was 
composed  of  portland  cement,  conglomerate  sand  and  trap 
rock  trommeled  over  %-in.  through  l^-in.  screens.  The  pro- 
portions used  were  1 :3  :5  in  wall  plates,  end  plates  and  divid- 
ings,  and  1:2:4  in  the  studdles  (or  struts).  The  reinforcement 
in  wall  and  end  plates  consisted  of  three,  %-in.  monolith-steel 
bars  with  ^-in.  webs  crimped  onto  them,  together  with  two 
straight  %-in.  bars.  The  dividings  were  reinforced  by  four 
i/^-in.  bars  wound  with  ^-in.  steel  wire,  the  whole  presenting 
a  column  with  square  cross-section.  Studdles  were  reinforced 
with  two  pieces  of  old  l^-in.  wire  rope.  Offsets  were  molded 
in  all  plates  5  in.  from  the  inside  face  to  accommodate  lining 
slabs.  The  design  of  the  different  members  is  clearly  shown 
in  Fig.  9. 

Reinforced-concrete  slabs  were  molded  for  the  shaft  lining, 
the  material  used  being  fines  of  trap  rock  (under  %  in.),  and 
conglomerate  sand,  with  Kahn  expanded  metal  as  reinforce- 
ment. The  mixture  used  for  slabs  was  1 :2 :4.  By  way  of 
experiment  a  piece  of  No.  1  hemlock  plank  of  the  same  length, 
width  and  thickness  as  a  concrete  slab  which  had  seasoned 
for  a  year  was  supported  at  either  end,  and  placed  side  by  side 
with  a  concrete  slab,  and  submitted  to  an  equal  pressure  applied 
across  the  center  of  each. 

Three  failure  cracks  appeared  in  the  concrete  slab  just 
previous  to  the  breaking  of  the  hemlock  plank,  although  total 


400 


COAL  MINING  COSTS 


collapse  of  the  concrete  slab  did  not  occur  until  the  pressure 
was  considerably  increased.  While  the  method  of  the  test 
employed  was  crude,  it  proved  that  the  concrete  slab  was  much 
superior  in  strength. 

A  concrete  mixer  of  the  drum  type  was  employed  in  pre- 
paring the  charge  for  the  forms.     The  amount  of  water  used 


v \ 


Long    Wall    Plate 
/      / 


I        k-$M 
Section 
A-B 


Short   Wall    Plate 
\       \      \      \ / 


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End    Piece 

4  Monolith  Bars  wound  with  kl'Web 


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X&BP 


Divider 


F7'^sWire 

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c^^w^^^^^^w^^r  pi^i^ 


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Lining    Slab 


FIG.  9. — Reinforced-concrete  sets  or  timbers  for  shaft  lining. 

in  the  mix  was  such  that  when  the  batch  was  piled,  it  settled 
rapidly  without  agitation.  A  drier  mix  was  attempted  by  way 
of  experiment,  but  owing  to  the  amount  of  reinforcement 
employed,  it  was  found  impossible  to  ram  this  concrete  into 
place. 

The  labor  force  required  was  six  men,  as  follows:  Two 
carpenters,  setting  up  forms  and  keeping  them  in  repair;  one 
man  wheeling  forms  onto  skidways  ready  for  filling,  return- 
ing used  forms  to  shop  and  cleaning  the  forms ;  one  man  feed- 


TIMBERING  COSTS  401 

ing  the  mixer  from  stick  piles  of  rock,  sand  and  cement;  one 
man  delivering  mix  to  forms  and  shoveling  material  into  place ; 
and  one  mason  ramming  charge  into  final  position.  With  this 
combination  of  men  as  many  as  four  complete  sets,  consisting 
of  64  separate  pieces,  have  been  molded  in  one  day  of  nine 
hours. 

The  shafts  lined  in  this  way  are  of  the  three-compartment 
type  (with  two  skipways  and  one  manway),  dipping  at  an 
angle  of  80  deg.  The  compartments  are  7  ft.  6  in.  high  inside, 
with  a  width  of  6  ft.  10  in.  for  the  skipways  and  3  ft.  for  the 
manways,  with  the  end  plates  and  dividings  making  the  great- 
est span  7  ft.  6  in. 

The  weights  of  the  different  pieces  comprising  the  set  are 
as  follows: 

Lb. 

Long  section  of  wall  plate 1035 

Short  section  of  wall  plate 700 

End  plate 600 

Dividers 645 

39-in.  studdles 268 

Complete  set,  16  pieces 8104 

Taking  the  weight  of  No.  1  Western  fir,  which  has  been 
exposed  to  the  weather  in  stock  piles,  as  33  Ib.  per  cu.ft.,  the 
above  concrete  set  weighs  almost  three  times  that  of  a  12  X 
12-in.  timber  set  which  the  concrete  set  is  intended  to  replace. 
Because  of  this  additional  weight  of  the  concrete  set,  it  was 
t'ound  necessary  to  increase  the  erecting  gang  from  the  usual 
5  or  6  men  on  the  timber  sets  to  7  men  for  the  concrete  sets. 
In  a  vertical  shaft  to  which  the  concrete  sets  are  especially 
adapted,  the  number  of  men  per  gang  might  be  reduced. 

The  comparative  cost  of  the  concrete  set  and  timber  set, 
delivered,  at  the  shaft  collar  is  striking.  The  concrete  set  was 
delivered  for  $22.50,  the  timber  set  for  $37.50.  These  figures 
are  based  on  the  following  prices : 

Western  fir $28.00  per  M.,  f.o.b.  car 

Crushed  rock 35c  per  yd.,  f.o.b.  shaft 

Conglomerate  sand 60c  per  yd.,  f.o.b.  shaft 

Portland  cement $1.15  per  bbl.,  f.o.b.  works 

Reinforcement $12.00  per  set,  f.o.b.  factory 


402  COAL  MINING  COSTS 

Cement  gun. — The  cement  gun  has  come  into  favor  for  cer- 
tain uses  in  connection  with  mine  timbering  and  the  following 
are  some  representative  costs  of  this  work  as  of  1920: 

In  the  examples  given,  details  of  cost  such  as  are  available 
have  been  presented,  and  are  incomplete,  in  that  power  and 
replacements  are  usually  omitted.  In  the  cement  gun  itself 
the  gaskets  for  the  cone  valves  require  replacement  from  time 
to  time,  as  does  the  outlet-valve  body  liner.  Liners  are  used 
for  the  nozzles.  Compressed  air  and  water  hose  are  subject  to 
the  wear  which  comes  from  frequent  handling,  and  would 
require  replacement  no  more  frequently  than  drill  hose.  The 
material  hose  is  subjected  to  considerable  wear.  Estimates  of 
its  life  range  from  four  to  six  months  with  continuous  use. 
Nozzle  liners  last  eight  days  and  cost  80c.  for  renewal.  The 
upkeep  cost  on  the  work  in  the  Anaconda  properties  at  Butte 
amount  to  50c.  per  eight-hour  shift  per  machine  operated. 

The  elements  of  cost  of  a  single  job  in  summary  are : 

1.  Assembly  of  machine,  compressed  air  and  water  piping, 
materials,  mixing  apparatus. 

2.  Preparatory  cleaning  of  surfaces,  placing  wire  reinforc- 
ing, staging. 

3.  "Guniting":    Labor,  power,  water,  cement,  sand,  lubri- 
cants. 

4.  Wear  and  replacement  of  liners,  gaskets,  hose,  gun  parts. 

5.  Tearing  down,  cleaning  up  and  removal  of  apparatus. 
H.  V.   Croll  has  given  the  cost  data  shown  in  the  table. 

In  this  example  "gunite"  was  used  to  prevent  the  walls  of  a 
mine  " tunnel"  from  slaking. 

M.  S.  Sloman  described  the  coating  of  a  coal  mine  slope 
with  gunite  in  1918.  The  surface  was  first  cleaned  and  scaled. 
No  reinforcement  was  used  and  the  coating  averaged  y2  in. 
in  thickness,  1 :3  mixture.  Timbers  were  covered  with  ^-in. 
wire  mesh.  The  slope  was  12  X  12  X  625°  ft.  The  total  cost 
was  $7488.58,  or  30c.  per  sq.  yd.  (3.3c.  per  sq.  ft.).  A  900-ft. 
section  of  this  slope  required  13!/2  8-hr,  shifts  for  a  working 
crew  of  eight  men ;  2376  sq.  ft.,  or  approximately  100  cu.ft., 
was  averaged  per  shift.  Materials  required  were  540  sacks 
cement  at  60c.  and  1620  sacks  sand  at  12c.  per  sack.  The 
working  crew  comprised  one  mechanic,  one  engineer  on  hoist, 
one  operator  on  cement  gun,  two  mixers,  one  nozzle  man,  one 


TIMBERING  COSTS  403 

man  drying  sand,  and  the  part  time  of  one  man  hauling  sand. 
No  mention  is  made  of  power  cost. 

COST  DATA  OF  PLACING  "GUNITE"  IN  TUNNEL  AT  UNITED  VERDE  EXTEN- 
SION MINING  Co.,  JEROME,  ARIZ. 

1  Gun  man,  also  motorman $  7 . 00 

1  Nozzleman 5 . 60 

1  Man  holding  lights 5 . 60 

1  Man  loading  gun 5 . 60 

1  Man  cleaning  roof 5 . 60 

3  Men  mixing  at  $5 . 60 16 . 80 


Total  labor .'.$  46.20 

Cement,  77  bags  at  $1 77 . 00 

Sand  9  cu.  yd.  at  $1 9 . 00 

Air  and  supplies 10 . 00 

Superintendence 5 . 00 

Total $147.20 

3750  sq.  ft.  at  $147.20  =  4c.  per  sq.  ft. 

Above  crew  placed  125  running  feet  of  tunnel  in  one  8-hour  shift,  equiv- 
alent to  3750  sq.  ft.;  average  thickness  |  in. 

Stephen  Royce  described  the  use  of  gunite  at  the  Gary 
"A"  shaft  at  Hurley,  Wis.  This  is  a  steel,  five-compartment 
shaft  with  the  steel  sets  blocked  in  place  with  wooden  block- 
ing and  lagging  of  3-in.  tamarack  planks  wedged  into  the 
flanges  of  the  I-beams.  "Gunite"  was  applied  to  fireproof  the 
lathing  and  wooden  blocking  and  to  protect  the  lagging  from 
decay  by  keeping  it  from  contact  with  air;  also  to  prevent 
corrosion  of  shaft  sets  and  water-proof  the  shaft.  The  surface 
to  be  covered  was  cleaned  thoroughly.  This  was  done  partly 
with  water  under  heavy  compressed  air  pressure,  partly  by 
sand  blasting,  and  partly  by  chipping  the  rust  and  accumulated 
coating  from  the  steel.  Reinforcement  consisted  of  No.  7, 
A.  S.  &  W.  Co.'s  triangular-mesh  reinforcing  wire  for  the  side 
walls.  This  was  stapled  directly  to  the  I-beams  and  to  the 
lagging  at  intervals.  To  keep  the  wire  mesh  about  one-eighth 
of  an  inch  away  from  the  surface  to  be  covered  was  ac- 
complished by  stapling  the  reinforcing  wire  on  with  nails 
underneath  it.  The  I-beams,  before  the  cement  was  applied, 
were  covered  with  1%-in.  mesh  chicken  wire,  clamped  on  with 
wire  clamps.  The  total  area  of  wall  surface  was  14,260.9 


404  COAL  MINING  COSTS 

sq.ft.;  of  steel  covered  3749.96  sq.ft.,  or  a  combined  total  of 
18,010  sq.ft.  The  materials  required  were  sand,  102.5  cu.yd. ; 
cement,  173  bbl. ;  reinforcing,  14,260.9  sq.ft.,  and  chicken  wire, 
3750  sq.ft.  Linear  feet  of  shaft  was  263.13.  The  work  required 
one  foreman  and  six  men  for  thirty-two  working  days.  The 
cost  was  given  as  $9.30  per  linear  foot  of  shaft.  As  the  area 
per  linear  foot  equaled  68.4  sq.ft.,  the  per-square-foot  cost  is 
found  by  calculation  to  be  13. 6c.  The  thickness  of  the  coat- 
ing was  given  as  1%  in.  The  "gunite"  was  applied  in  from 
one  to  three  coats. 

Reclaiming  timbers. — Although  mine  props  and  sets  of  tim- 
ber are  often  broken  a  short  time  after  being  set,  the  broken 
ends  are  valuable  as  they  can  still  be  utilized  for  the  purpose 
of  cap-pieces,  wedges,  track  ties,  or  for  the  building  of  "cogs" 
or  "chocks."  Also,  post  timber  broken  in  a  thick  seam  can 
often  be  used  again,  in  a  thinner  seam  at  the  same  colliery. 

In  some  mines  there  is  a  considerable  loss  of  timber,  through 
the  carelessness  of  miners  who  will  let  them  lie  in  the  waste 
where  they  are  finally  buried.  By  keeping  a  careful  watch  in 
their  daily  rounds  through  the  mine,  the  mine  officials  can  do 
much  toward  reducing  this  loss  or  waste  of  timber. 

It  is  well  to  emphasize  the  importance  of  drawing  all  kinds 
of  timber,  as  the  work  proceeds,  using,  if  necessary,  some 
suitable  appliance  for  this  purpose.  Timbers  left  standing  in 
the  waste  often  cause  a  loss  greater  than  their  own  value,  by 
preventing  the  roof  from  caving  and  frequently  making  it 
necessary  to  build  extra  packwalls  or  timber  cogs  to  keep  the 
roads  open.  The  material  for  these  packwalls  often  has  to  be 
transported  a  considerable  distance;  whereas,  if  the  timber 
was  drawn  and  the  roof  allowed  to  fall,  there  would  be  plenty 
of  material  for  the  building  of  all  necessary  packwalls  in  most 
cases. 

Again,  under  many  conditions,  when  the  roof  does  not  fall 
but  a  large  standing  area  is  kept  open  a  great  weight  is  thrown 
on  the  timbers  standing  next  to  the  face  of  the  coal,  with 
the  result  that  these  timbers  are  broken  more  quickly,  or  they 
kick  out  and  the  roof  is  ruptured  at  the  face.  When  this 
occurs,  the  condition  is  bad,  as  the  influence  of  the  roof  in 
breaking  the  coal  after  the  latter  is  undermined,  is  destroyed. 
When  the  roof  is  of  such  a  nature  that  it  breaks  readily,  it 


TIMBERING  COSTS  405 

is  a  very  good  policy  to  set  a  line  of  large  breaking  posts, 
with  good  sized  cap-pieces,  on  one  side  of  the  track,  which 
should  be  carried  along  the  straight  rib  of  the  room. 

As  the  face  of  the  room  advances,  up  to  the  last 
crosscut,  there  is  not  only  a  saving  of  timber,  but  the  caving 
of  the  roof  prevents  the  crushing  of  the  pillar  coal  when  the 
roof  "weights"  and  cannot  fall.  In  heavy  pitching  seams,  the 
recovery  of  timbers  is  much  more  difficult  and  dangerous  than 
in  flat  seams ;  because  the  worked-out  portion,  from  which  the 
timbers  are  drawn,  is  located  up  the  pitch,  and  any  loose 
pieces  of  rock  that  fall  when  the  post  is  drawn  are  liable  to 
roll  or  slide  down  upon  the  men  engaged  in  drawing  the  timber 
who  are  unprotected. 

The  danger  may  be  avoided,  in  part  or  wholly,  by  using 
a  long  i^-in.  steel  cable  or  chain  that  will  reach  from  the  tim- 
bers to  the  first  crosscut,  in  which  the  drawing  machine  should 
be  placed.  This  will  not  only  afford  the  necessary  protection 
for  the  men,  but  will  enable  them  to  recover  a  larger  percentage 
of  timbers.  The  cost  of  timber  in  pitching  seams  is  much 
greater  than  in  flat  seams,  owing  to  the  labor  required  in 
handling  the  timber  on  steep  pitches.  In  a  seam  pitching  35 
deg.  the  cost  of  timber  frequently  amounts  to  an  average  of 
about  S^c.  per  ton  of  coal  mined.  This  was  in  a  mine  where 
the  roof  conditions  were  fairly  good. 


V 
MISCELLANEOUS  INSIDE  COSTS 

TUNNELING  COSTS 

American  and  foreign  tunneling  records  compared. — The 
accompany  table  gives  some  of  the  most  creditable  American 
tunnel  records  made  up  to  1909.  The  ranking  order  of  some  of 
the  tunnels  is  probably  open  to  discussion.  That  given,  based 
on  the  31-day  record,  is  by  no  means  necessarily  the  order  of 
merit.  There  is  a  tendency  among  mining  men  to  mention 
the  highest  one  month's  record,  however,  rather  than  to  recall 
record  figures  extending  over  longer  periods. 

A  comparison  of  foreign  records  with  our  own  is  shown 
in  another  table.  On  the  face  of  this  comparison,  American 
records  appear  to  rank  ahead  of  the  Continental  records.  Thus 
the  best  month's  record  on  the  Gunnison  tunnel  actually 
exceeds  the  best  Simplon  record  by  87  ft.  While  we  have  no 
explicit  data  concerning  the  kind  of  rock  in  which  the  record 
progress  for  the  Simplon  was  made,  we  know  that  the  Gun- 
nison record  was  made  using  air-driven  augers  in  the  soft  shale 
of  the  east  heading,  in  which  7500  ft.  of  progress  was  made 
for  the  first  year,  making  the  high  average  of  625  ft.  per  month. 
In  the  granite  of  the  west  heading,  the  best  Gunnison  record 
449  ft.  per  month  so  that  this  is  the  figure  that  should  in  all 
fairness  be  compared  with  the  Simplon  record  of  755  ft.  While 
no  long-time  average  figure  is  available  for  the  Gunnison 
progress  in  granite,  we  can  hardly  anticipate  a  serious  rival 
to  that  remarkable  record  in  the  Simplon  of  426  ft.  per  month 
for  76  months.  The  top  notch  American  record  in  granite, 
however,  was  that  made  in  October,  1908,  by  the  Elizabeth 
Lake  tunnel  at  Los  Angeles,  Calif.,  where  a  progress  of  466  ft.. 
was  made  in  the  31  days  ending  Oct.  31. 

406 


MISCELLANEOUS  INSIDE  COSTS 


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408 


COAL  MINING  COSTS 


COMPARISON  OF  AMERICAN  AND  FOREIGN  RECORDS 


Name 

Continental 

American 

Simplon  Tunnel 

Gunnison  Tunnel 

1 

Best  month's  record 
Best  long-time  record 

755  ft. 
426  ft.  per  month  for  76  months 

842  ft.  shale,  449  ft.  granite 
625  ft.  for  12  months  (shale) 

Name 

St.  Gothard  Tunnel 

Roosevelt  Tunnel 

2 

Best  month's  record 
Best-long  time  record 

563  ft. 
263  ft.  per  month  for  93  months 

435  ft. 
292  ft.  for  12  months 

A  mere  comparison  of  records,  however,  is  more  favorable 
to  Americans  than  will  appear  because  of  the  difficulties  of 
driving  the  Continental  tunnels  due  to  great  length  and  depth 
beneath  the  surface,  resulting  in  excessive  rock  pressure  and 
high  temperature,  not  to  mention  copious  flows  of  hot  water. 
In  none  of  our  long  American  tunnels,  unless  it  was  the  Corn- 
stock  at  Virginia  City,  Nev.,  were  conditions  of  high  tempera- 
ture and  flows  of  hot  water  encountered,  comparable  to  those 
of  the  Simplon  and  St.  Gothard  tunnels.  For  the  Sutro,  by 
the  way,  the  best  month's  progress  is  recorded  as  417  ft. 

In  case  it  may  be  argued  that  it  is  unfair  to  compare  rail- 
way tunnels  with  mine  tunnels,  it  is  only  necessary  to  point 
out  that  all  data  given  relate  to  the  speed  of  driving  the  head- 
ings only,  whose  size  does  not  greatly  differ  in  a  railway  tunnel 
from  a  mine  adit. 

The  argument  often  advanced  in  excuse  of  American  back- 
wardness is  that  the  breaking  qualities  of  the  rocks  penetrated 
by  our  tunnels  are  much  inferior  to  those  of  the  Continental 
tununels.  In  the  accompanying  tables  the  seventh  column 
gives  the  type  of  rocks  encountered.  This  will  not  help  the 
comparison  much  except  in  a  general  way,  for  every  mining 
man  knows  that  what  may  be  a  soft  rock  in  one  locality  is 
hard  and  tough  in  another.  Hardness  of  rock  is  hardly  a  fair 
basis  of  comparison,  nor  is,  indeed,  specific  gravity.  It  is  a 
well-known  fact  that  in  many  of  our  American  tunnels,  some 
of  the  hardest  rocks  to  drill  actually  blast  the  best,  while  the 
best  Simplon  tunnel  records  are  said  to  have  been  made  in  the 
hard  Antigorio  gneiss. 


MISCELLANEOUS  INSIDE  COSTS 


409 


a 
o  a 


1 1 

sw 


§s 

2  ^ 

si 


Long-Ti 
Average 


ft.  per  month 
76  months 


s, 
ne 
h 


Jll 


11*1 

li'iij 

•SJ1-  SI 
s  s  s  s  s  a 


a  8 


0 


co" 


0  13 

o>    s 


§      Si 

•**  *'3     c8 

I  8,1-5 


a      s 
15-8 

^  -s 
•<     > 

OQ 


« 


OQ 


410  COAL  MINING  COSTS 

In  America,  the  heavy  Sullivan  drills  made  the  first  Ophelia 
tunnel  record  at  Cripple  Creek,  and  then  beat  it  by  54  ft.  on 
the  west  heading  of  the  Gunnison.  The  Leyner  No.  9  hammer 
drill  beat  the  Ophelia  record  at  the  Roosevelt  tunnel  work,  and 
finally  captured  the  American  record  previous  to  1910  by  its  run 
of  466  ft.  at  the  Elizabeth  Lake  tunnel,  Los  Angeles,  Calif., 
October,  1908. 

In  Europe  the  Ferroux  percussion  drill  and  the  Brandt 
rotary  core  drill,  both  of  foreign  make,  have  established  a  long 
series  of  records  that  threw  the  performance  of  American 
drills  far  in  the  rear  until  the  appearance  of  the  Ingersoll- 
Rand  drill  at  the  Loetschberg  tunnel  in  1906.  Since  then,  this 
American  drill  has  not  only  eclipsed  all  American  records  by 
its  September,  1907,  record  progress  of  574  ft.,  but  has  exceeded 
all  the  previously  established  foreign  records  except  three, 
namely,  the  Simplon  (755  ft.),  the  Arlberg  (641  ft.),  and  the 
Albula  (607  ft.). 

But,  if  any  one  individual  machine  is  to  have  the  credit  for 
the  world  record  performance  of  tunnel  driving  up  to  1909,  that 
credit  belongs  to  the  Jeffrey  A-2,  air-driven  auger  with  which 
842  ft.  of  soft  shale  were  removed  in  one  month's  time  from 
the  east  heading  of  the  Gunnison  tunnel. 

When  the  Brandt  drill  accomplished  the  wonderful  records 
of  the  Simplon,  some  engineers  went  so  far  as  to  say  that  its 
success  spelled  the  finish  of  air-driven  percussion  drills.  Others, 
however,  pointed  to  the  fact  that  in  driving  the  Arlberg  tunnel 
with  Ferroux  percussion  drills  in  one  heading  and  with  the 
Brandt  rotary  drills  in  the  other,  the  best  month's  record  of 
the  two  differed  by  less  than  1  per  cent.  As  a  matter  of  fact, 
an  average  pf  the  four  best  months  *  records  of  the  Ferroux 
drills  was  613  ft.  as  against  576  ft.  of  the  Brandt  drills.  The 
top-notch  record  of  each  was  637  ft.  for  the  Ferroux  and  641 
ft.  for  the  Brandt. 

The  secret  of  the  great  speed  made  in  the  Alpine  tunnels 
appears  to  have  been  in  the  very  careful  study  made  of  the 
various  causes  of  delay  in  the  successive  operations  of  drilling, 
blasting,  mucking  out,  and  setting  up  the  drills  again.  As 
a  result,  a  radically  different  method  of  mounting  the  drills 
in  the  heading  has  been  employed  from  that  practiced  in 


MISCELLANEOUS  INSIDE  COSTS  411 

America.  In  the  mounting  of  the  drills,  the  return  to  the  old 
carriage  drill  is  seen,  but  there  is  no  such  cumbersome  affair 
as  the  drill  carriage  used  in  the  early  tunneling  operations  in 
this  country. 

The  drill  carriage  at  the  Loetschberg  consists  simply  of  a 
small  truck  whose  wheel  base  is  about  4  ft.,  running  on  the 
regular  heading  track.  Mounted  on  this  truck  in  the  longi- 
tudinal axis  of  the  tunnel,  and  hinged  to  swing  vertically,  is 
an  I-beam  set  with  its  web  vertical  and  reinforced  to  give  it 
lateral  stiffness.  On  the  forward  end  of  the  I-beam  there  is 
mounted  what  would,  in  America,  be  called  a  shaft  bar,  set 
transversely  while  the  beam  is  pivoted  so  as  to  swing  the 
bar  horizontally.  On  the  opposite  end  of  the  I-beam  is  mounted 
a  counterweight.  Four  drills  are  mounted  on  the  shaft  bar, 
the  compressed-air  connections  from  these  drills  running  back 
to  one  hose  connection  in  the  rear  of  the  truck.  When  not  in 
use,  the  shaft  bar  is  swung  so  that  it  lies  directly  over  the 
I-beam  and  the  carriage  can  then  be  run  anywhere  over  the 
heading  tracks  occupying  no  more  room  in  the  heading  than 
two  muck  cars.  This  carriage  is  practically  the  same  as  that 
used  by  Brandt  on  the  Simplon  tunnel,  the  main  difference 
between  the  two  systems  of  work  being  found  in  the  drills. 

Before  blasting,  a  %-inch  plate  of  steel  about  6%  X  3%  ft. 
is  laid  down  just  ahead  of  the  track  end.  After  blasting,  the 
gases  are  sucked  out  of  the  heading  through  a  pipe  of  24  in. 
diameter,  the  fan  being  capable  of  running  either  as  an  ex- 
hauster or  blower.  A  cut  is  then  mucked  through  the  center 
of  the  muck  pile  down  to  the  steel  plate  sufficiently  wide  to 
allow  the  arm  of  the  drill  carriage  to  introduce  the  bar  carry- 
ing the  drills  into  the  top  01  the  heading,  the  drill  carriage, 
of  course,  running  on  the  steel  plate  laid  down  ahead  of  the 
track.  The  shaft  bar  is  then  jacked  firmly  against  the  sides  of 
the  heading  and  drilling  immediately  begins  on  the  top  holes. 
Mucking  out  then  continues  while  drilling  is  in  progress.  On 
account  of  the  drills  remaining  on  the  carriage  all  the  time  with 
the  air  connections  at  the  drills  undisturbed,  there  is  little 
chance  for  grit  to  get  into  the  working  parts  and  so  impair 
their  efficiency.  The  following  table  gives  the  approximate 
time  required  for  the  various  operations  in  the  heading. 


412  COAL  MINING  COSTS 

Minutes 

Setting  up  drill  carriage  in  the  heading 20 

Drilling  12-14  holes,  4  ft.  deep,  2  in.  diameter,  at  bottom .  60 

Removing  drill  carriage  from  heading 20 

Loading  and  firing  holes 30 

Clearing  out  smoke 20 

Cutting  center  of  muck  pile  to  admit  carriage 90 

240 

=  4  hours 

As  before  mentioned,  work  is  carried  on  in  three  8-hr, 
shifts,  each  shift  being  expected  to  drill  two  rounds  and  shoot 
twice  making  7  ft.  per  shift  advance  or  21  ft.  per  day. 

Here  we  have  an  American  drill  adapted  to  a  European 
system  of  work,  beating  all  American  records  and  threatening 
to  rival  those  of  the  Simplon  before  construction  is  completed. 
It  is  apparent  that  system  rather  than  the  various  European 
makes  of  drills  is  to  be  credited  with  the  Continental  records. 

It  is  interesting  to  compare  these  time  items  with  those 
of  the  Simplon  tunnel  which  were  described  in  a  paper  on 
tunneling,  read  before  the  Eoyal  Institute  of  Great  Britain, 
May  5,  1900: 


Bringing  up  and  adjusting  the  dri 
Drilling                                   

Minutes 
11....          20 
105 

Minutes 
25 
150 

Charging  and  firing 

15 

15 

Cleaning  UD  rock  debris  .  . 

120 

120 

Total, 


260  305 

=  4  hours,       =5  hours 
20  minutes     5  minutes 


The  average  advance  made  by  this  method  is  given  as  3  ft. 
9  in.,  or  about  7%  ft,  per  8-10-hr  shift. 

Perhaps  the  first  American  engineer  to  successfully  apply 
a  bonus  system  of  payment  for  tunnel  driving  in  addition  to 
the  usual  wages  was  D.  W.  Brunton,  of  Denver.  The  follow- 
ing were  the  rates  of  payment  used  by  him  in  driving  the 
Cowenhoven  tunnel  at  Aspen,  Colo.,  in  1889,  when  the  progress 
made  per  month  exceeded  150  ft. 


MISCELLANEOUS  INSIDE  COSTS  413 

Bonus  for 
Progress,  Feet  every  Foot 

150-200 $1.00 

200-250 1.50 

250-300 2.00 

300-350 2.50 

Over  350 3.00 

With  these  payments  drillmen  often  made  $120;  helpers, 
$112 ;  and  laborers,  $95  per  month  in  addition  to  their  regular 
wages. 

Wherever  tried  in  America,  the  bonus  system  has  generally 
proved  a  success.  This  is  for  the  simple  reason  that  an  unruly 
giant  will  not  especially  exert  himself  to  earn  a  $3  or  $4  per 
shift.  But  make  him  a  bonus-system  proposition  whereby 
exertion  of  wits  as  well  as  muscles  may  bring  him  from  $6  to 
$8  per  shift,  and  we  have  a  transformation  from  halfway  in- 
difference to  keen  interest  in  his  work.  Stimulating  the  spirit 
of  competition  between  the  various  crews  has  in  some  cases 
given  good  results  in  increased  work  but  in  no  case  has  the 
glory  of  beating  the  other  fellow  proved  so  satisfying  as  that 
extra  $2  or  $3  per  shift. 

Cost  of  rock  tunnel  at  a  coal  operation. — As  a  general  rule 
the  only  literature  on  driving  rock  tunnels  is  that  descriptive 
of  work  in  metal  mines,  or  for  irrigation,  water  supply  and 
railroads.  Such  tunnels  are  usually  driven  by  organizations 
specializing  in  this  work  and  much  effort  is  expended  toward 
making  records. 

In  coal-mining  operations  and  the  opening  of  new  coal  fields 
it  is  often  necessary  to  drive  rock  tunnels  but  only  small  men- 
tioin  is  made  of  them  and  the  special  methods  used  and  equip- 
ment available. 

The  tunnels  here  described  were  used  to  open  a  new  develop- 
ment adjacent  to  the  Utah  Fuel  Co.'s  Clear  Creek  mine  No.  1, 
about  1914,  and  were  driven  as  haulage  and  air-course  entries. 

In  order  to  systematize  and  expedite  the  work  as  much 
as  possible  a  separate  organization  was  created — made  up  of 
expert  hard-rock  men.  Although  kept  distinct  from  the  regu- 
lar operating  organization  it  was  necessary  to  coordinate  their 
work  with  that  of  the  mine  in  order  not  to  interfere  with  the 
production  of  coal. 


414  COAL  MINING  COSTS 

The  main  haulage  rock  tunnel  was  driven  with  a  rectangu- 
lar cross-section  having  a  minimum  height  of  8  ft.  and  a 
minimum  width  of  10  ft.,  while  the  air  course,  of  a  similar 
cross-section,  had  minimum  dimensions  of  8  ft.  in  height  and 
12  ft.  in  width.  Over  breakage  and  trimming  have  increased 
these  dimensions  on  an  average  of  one  foot  each  way.  Two 
tunnels  were  driven  in  order  to  provide  for  adequate  ventila- 
tion of  the  new  mine  as  it  was  impracticable  to  sink  a  shaft 
or  drive  an  adit  from  the  outside  for  air. 

The  rock  strata  penetrated  were  sandstones  and  shales  of 
medium  hardness.  The  shales  were  the  hardest  to  drill  as  the 
gumming  of  the  cuttings  made  it  difficult  to  keep  the  holes 
clean.  The  first  halves  of  the  tunnels,  driven  on  a  y2  per  cent 
grade,  were  parallel  with  the  bedding  planes  of  the  strata  and 
thus  much  harder  to  break.  The  last  halves,  driven  on  a  7^/2 
per  cent  grade,  took  the  tunnels  off  the  bedding  planes  some- 
what but  increased  difficulties  were  encountered  in  the  numer- 
ous small  faults  which  were  cut  through  as  a  main  faulting 
plane  was  approached. 

On  the  average  the  ground  was  fair  drilling  but  in  some 
places  it  was  necessary  to  use  water  under  pressure  to  clean 
out  holes  so  as  to  render  it  possible  to  drill  them  at  all.  In 
such  cases  the  water,  at  100  Ib.  pressure,  was  forced  into  the 
holes  through  %-in.  pipe  16  in.  long  used  as  a  nozzle.  The 
time  necessary  to  drill  the  rounds  varied  from  one  hour  and 
45  min.  to  the  full  8-hr,  shift.  The  outside  back  holes  were 
drilled  first,  then  the  outside  breast  holes,  then  the  inside  breast 
holes,  and  so  on  as  shown  in  the  accompanying  sketches,  Fig. 
1,  the  numbers  designating  the  order  of  drilling. 

The  source  of  power  was  the  power  plant  of  the  mine,  con- 
sisting of  six  125-hp.  72-in.  by  16-ft.  horizontal  return-tubular 
boilers  hand  fired  with  slack  or  the  poorest  grade  of  run-of- 
mine  coal  coming  from  the  mine,  and  two  Thompson-Ryan 
generators  of  175-kw.  capacity,  320  amp.  500  volts,  driven  by 
19  X  18-in.  McEwan  engines  of  260  hp.  each  at  200  r.p.m. 
Power  was  charged  against  the  work  at  the  rate  of  1.24c.  per 
kilowatt-hour. 

Current  from  the  generators  was  conducted  about  200  ft. 
to  a  24  X  64  ft.  combination  compressor  and  blacksmith  shop, 
40  longitudinal  feet  of  which  was  used  for  the  compressors. 


MISCELLANEOUS  INSIDE  COSTS 


415 


Two  Ingersoll-Rand  compressors  of  the  Imperial  type  No.  10, 
were  used.  These  compressors  had  a  rating  of  427  cu.ft.  of 
free  air  per  minute  at  175  r.p.m.  at  sea  level.  Each  machine 
was  driven  through  a  belt  connection  from  a  500-volt  116-amp. 
75-hp.  General  Electric  continuous-current  motor  operating  at 
850  r.p.m. 


B-^ec  oncf position  4?  bar; 
double  shi ft. 
SIDE   ELEVATION 

Holes  large  so  charge  could 
be  placed  welf  toward  back 
of  holes 

[>ouble  primers  used  in  Lifter 
and  Cut  holes,  also  m  wethotes\ 

'Fuse  cut  2'difference  for 
'different  kinds  of  holes, 
^"difference  tor  Cut  holes. 


HOLE  DATA  16  HOLE  ROUND 


KIND  OF 
HOLE 

LENGTH 
INSIDE   RD. 

LENGTH 
OUTSIDE  RD. 

STICKS  OF  60% 
POWDER  8xlV 

Back 

6(-4v 

_5-6" 

8  to  10 

Breast 

6-8" 

5-10' 

9  toll 

Cut 

7-  Z" 

6V  4* 

12  to  15 

Lifters 

6-4" 

5'-  6" 

IZtolB 

r 

STEEL 

DATA 

KIND 

MATERIAL 

LENGTH 

sire: 

GAGE 

Starters 

Cruciform  Steel 

Z'-6" 

Z' 

3" 

Seconds 

CruafbrmStieel 

4-6" 

iV 

2&* 

Thirds 

Cruciform  Steel 

6-6" 

I  YZ' 

2!^" 

Fourths 

Cruciform  Steel 

!L&"-&'-6 

[£ 

2i4r 

FIG.  1. — Method  of  drilling  rock  tunnel  at  the  Sunnyside  Mine. 

Air  was  conducted  to  two  receivers  in  the  compressor  house, 
one  being  3  ft.  in  diameter  by  8  ft.  3  in.  long  and  the  other 
3  ft.  6  in.  by  8  ft.,  then  into  the  mine  through  a  4-in.  pipe 
3725  ft.  long  to  a  3-ft.  6-in.  by  8-ft.  receiver,  thence  350  ft. 
through  a  3-in.  pipe  line  to  the  starting  points  of  the  tunnels 
and  through  3-in.  branch  lines  to  within  50  ft.  of  the  faces, 
1-in.  air  hose  being  used  from  thence  to  the  drills. 

Drill  steel  was  sharpened  with  a  Numa  rock-drill  sharpener. 
For  heavy  blacksmith  work  the  smith  set  up  an  old  Sullivan 


416  COAL  MINING  COSTS 

coal  puncher  for  a  hammer.  This  required  only  the  fitting  on 
of  a  hammer-block  4  X  6  X  6  in.  in  place  of  the  bit  of  the 
puncher,  setting  the  machine  on  a  suitable  frame  and  making 
a  foot  control.  The  expense  of  this  makeshift  was  slight  but 
the  saving  effected  in  the  cost  and  in  the  grade  of  work  done 
was  quite  noticeable.  For  forge  fires  a  combination  of  coke 
breeze  and  slack  coal  from  the  Sunnyside  mines  of  the  com- 
pany gave  good  results. 

Chicago  giant  rock  drills,  size,  3%  in.  were  used  in  driving, 
while  jackhamers  and  Sullivan  stoping  drills  were  used  for 
trimming  and  widening.  Two  drills  were  used  at  each  face 
and  these  were  mounted  on  7-ft.  double  screw  columns.  On 
double  shift  work,  the  muck  was  always  in  the  way  during  the 
first  drilling  so  it  was  necessary  to  mount  the  drills  on  10%-ft. 
single  screw  bars,  having  a  single  screw  brace  to  the  face  to 
take  up  the  vibration  in  the  bar.  In  this  way  the  upper  rows 
of  holes  were  drilled  first  with  the  bar  set  near  the  roof  and 
by  the  time  these  were  completed  the  muck  was  cleared  away 
from  the  face  and  the  bar  could  be  set  up  in  a  lower  position 
and  the  remainder  of  the  holes  drilled.  A  3-in.  water  main 
having  %-in.  hose  connections  was  used  along  the  back  entry 
to  within  50  ft.  of  the  face.  From  these  connections  to  the 
faces,  %--in.  pipe  was  used  with  %  X  16-in.  special  reducing 
nozzles. 

The  explosive  used  was  mainly  60  per  cent  dynamite  with 
sticks  8  in.  long  and  l1/^  in.  in  diameter. 

On  single  shift  the  machine  men  worked  during  the  day, 
that  is  from  7  :30  a.m.  to  4  p.m.,  and  the  muckers  and  drivers 
during  the  night  shift  beginning  at  8  p.m.  and  ending  when 
the  muck  was  all  cleaned  up.  The  shift  for  the  machine  men, 
muckers  and  drivers,  while  nominally  8  hr.,  was  considered 
finished  whenever  their  work  was  completed.  A  bonus  was 
given  of  10  per  cent  of  the  day's  wage  for  each  foot  over  3  ft. 
of  tunnel  driven  that  they  averaged  per  shift.  The  bonus  was 
calculated  at  the  end  of  the  month  instead  of  each  day,  except 
in  the  case  of  men  quitting  before  the  end  of  the  month.  The 
average  wages  earned  by  this  system  by  the  machine  men  was 
$4.17  per  shift.  These  men  were  required  to  do  the  drilling, 
extend  both  the  air  and  water  lines,  make  up  primers,  load 
and  shoot  holes. 


MISCELLANEOUS  INSIDE  COSTS  417 

Timbermen  were  paid  the  same  wages  as  machine  men  in- 
cluding the  bonus.  There  was  comparatively  little  timbering 
to  be  done  so  that  machine  men  were  used  for  this  work  and 
it  was  thought  fair  to  allow  them  the  same  wages  as  they  would 
have  earned  if  running  machines. 

There  were  four  muckers  to  each  heading  who  were  paid 
$3.25  per  8-hr,  shift  and  a  bonus  of  10  per  cent  of  the  day's 
wage  for  each  foot  over  three  feet  of  tunnel  driven  per  shift, 
just  as  paid  the  machine  men.  By  this  system  the  average 
wage  earned  by  muckers  was  $3.84  per  shift.  The  muckers  laid 
all  track,  not  including  switches,  put  down  the  mucking  sheets, 
loaded  and  unloaded  drill  steel  and  kept  track  clean  for  100  ft. 
from  the  faces. 

Drivers  were  paid  $3.15  per  8-hr,  shift  without  any  bonus. 
They  were  required  to  take  muck  away  from  the  faces  to  the 
parting,  bring  in  empties  and  haul  steel,  powder,  caps,  fuse, 
rails,  ties,  etc.,  from  parting  to  faces. 

The  dumpers  outside  were  paid  $3  and  $3.15  per  8-hr,  shift 
without  any  bonus.  They  were  required  to  handle  the  rock 
but  the  extending  of  the  tipples  was  done  by  the  regular  mine 
carpenters  paid  $3.40  and  $3.15  per  8-hr,  shift. 

The  blacksmith  was  paid  $4  per  8-hr,  shift  and  straight 
time  for  overtime.  His  duties  were  the  sharpening  of  drill 
steel,  general  repairs  and  other  blacksmith  work.  The  com- 
pressor attendants  were  paid  $3.50  per  8-hr,  shift  and  were 
required  to  run  the  compressors  and  assist  the  blacksmith.  The 
head  foreman  was  paid  $175  per  month  and  his  assistant  $4.50 
per  8-hr,  shift. 

The  work  cf  widening  the  main  haulage  tunnel  for  the  pass- 
by  and  parting  was  done  after  the  tunnels  were  driven  to  the 
coal.  For  this  work  Ingersoll-Rand  Jackhamer  drills  with  %-m- 
hollow  steel  were  used.  The  men  were  paid  for  straight  time 
at  the  same  rate  as  for  tunnel  driving,  no  bonus  being  given. 
The  shots  were  detonated  by  electricity,  using  6X  detonators. 
A  total  lineal  distance  of  700  ft.  8  ft.  high  was  widened  from 
3  to  5  ft.,  the  unit  cost  for  the  work  being  given  on  p.  418. 

If  the  cost  of  the  widening  operations  is  added  to  that  of 
driving  the  tunnels  the  total  expense  per  foot  would  be  about 
$17,  but  if  salvage  is  allowed  on  pipe  and  material  left  over 


418 


COAL  MINING  COSTS 


UNIT  COSTS  OF  STRAIGHT  TUNNEL 

Engineering $0. 0503 

Superintendence 0 . 9885 

.,,.  f  Machine  men 

lmg \  Machine  men  bonus 

Muckers 

Muckers'  bonus 

Dumpers 

Extending  dump 

Drivers 

Motor  men 

Shafts 

Stable  expense 

Horse  killed 

Ditch  and  track 

Tracks  and  switches 
Blacksmithing 0.3844 

P.H.  charge 

Running  compressor 


Handling  rock 


Hauling . 


3 . 8221 


4.0331 


1.3193 


Power.  . 


1.8604 


Repairs,  compressor 

Pipe  lines 

Repairs,  drills 

Timbering 0. 4389 

Blasting  material 2. 7707 

Housing 

Setting 
Machinery...       Depreciation 

Miscellaneous 
Total  cost  per  foot  of  tunnels .  .  . 


0.6655 


$16.3332 


COST  PER  LINEAL  FOOT  OF  WIDENING  OPERATIONS 

Superintendence $0. 4675 

Drilling 0. 3656 

Muckers 

Handling  rock  \  Dumpers                      J.  1 . 3243 

Extend  dump 

Drivers 

Hauling \  Stable  expense             }•  0. 3171 

Motor  men 

Blacksmithing 0. 1712 

(  Compressor  men          \  ., 

Power (Pipelines                     }  °'1681 

Blasting  material 0. 5292 

Miscellaneous. .                                          0 . 0486 


Total  cost  per  linear  foot . 


$3.3916 


MISCELLANEOUS  INSIDE  COSTS  419 

and  subsequently  used  for  coal-mining  operations  the  cost  per 
foot  is  brought  down  to  $15,  which  would  be  the  net  cost. 

The  average  progress  per  shift  for  both  single-  and  double- 
shift  work  during  the  whole  time  of  driving  the  tunnels  was 
5.11  ft.  The  highest  average  per  shift  per  month  was  5.56  ft. 
The  greatest  distance  driven  in  a  tunnel  in  a  single  month  was 
323  ft.  in  60  shifts. 

Single-shift  work  was  found  to  be  less  efficient  than  double- 
shift  work,  although  to  a  certain  extent  this  was  due  to  the 
forming  of  a  working  organization  and  to  the  necessary  experi- 
menting with  the  drilling  and  shooting  of  the  ground  during 
the  earlier  stages  of  the  undertaking. 

Some  examples  of  tunnel  costs. — A  tunnel  located  at  Idaho. 
Springs,  37  miles  west  of  Denver,  in  the  lower  Clear  Creek 
mining  district  of  Colorado  started  in  1893  and  extended  at 
intervals  over  a  10-yr.  period,  presents  some  interesting  cost 
data.  The  original  purpose  of  this  tunnel  was  to  cut  the  well- 
known  veins  on  the  line  of  the  tunnel  so  as  to  receive  royalties 
from  the  property  owners  for  drainage  and  for  transportation 
of  ore. 

The  yearly  progress  made  is  as  follows :  1893,  80  ft. ;  1894, 
1405  ft. ;  1895,  1992  ft. ;  1896,  2061  ft. ;  1897,  1080.5  ft. ;  1898, 
99  ft;  1899,  455  ft.;  1900,  2285.3  ft.;  1901,  2923.5  ft.;  1902, 
761.1  ft. ;  1904,  1389.3  ft. ;  1905,  672.5  ft.  A  summing  up  of  the 
above  footages  gives  a  total  length  of  15,154.7  ft.  from  the 
original  starting  point. 

The  first  80  ft.  of  the  tunnel  were  driven  by  hand,  but  as 
there  were  no  available  data  concerning  this  hand  work,  very 
little  can  be  said  about  it.  This  method  necessarily  made  the 
progress  of  the  work  very  slow,  and  it  was  deemed  advisable 
to  install  a  compressed-air  plant  to  supply  power  drills.  The 
equipment  was  in  duplicate,  consisting  of  two  Norwalk  com- 
pound 14  X 16  in.  high-altitude  compressors  and  two  80-hp. 
boilers.  These  compressors  supplied  air  to  the  drills,  which 
were  of  the  3-in.  Leyner  percussion  type.  From  100  Ib.  to 
175  Ib.  of  dynamite  were  used  to  the  round,  and  the  greatest 
footage  made  was  160  ft.  per  month  by  a  working  force  of  26 
men. 

In  October,  1899,  there  was  a  change  in  the  management, 
and  the  work  was  done  in  a  more  systematic  manner.  The 


420  COAL  MINING  COSTS 

equipment  of  the  power  plant  was  increased  by  the  addition  of 
another  80-hp.  boiler  and  a  22  X  24  in.  compound  Norwalk 
compressor. 

The  holes  were  placed  in  the  breast  according  to  the  Ameri- 
can center-cut  system  with  an  extra  plunger  hole  at  the  upper 
center  of  the  cut.  The  side  and  cut  holes  were  drilled  by  two 
3-in.  Leyner  sluggers  mounted  directly  on  separate  columns 
placed  on  each  side  of  the  tunnel,  while  one  model  5  Water 
Leyner,  mounted  on  an  arm,  drilled  the  back  and  plunger  holes. 
One  Slugger  was  started  at  the  bottom  and  worked  up,  while 
the  Water  Leyner  put  in  the  back  holes  on  that -side.  The 
other  Sluggers  started  at  the  top  and  worked  down  so  that 
it  was  out  of  the  way  when  the  Water  Leyner  machine  was 
ready  to  shift  to  the  other  column. 

This  system,  together  with  the  use  of  high  pressure  air,  160 
lb.,  resulted  in  deeper  holes  in  less  time,  and  therefore  greater 
progress.  In  blasting,  the  cut  holes  were  electrically  fired  first, 
then  others  until  the  whole  cut  was  taken  out  clean,  using  about 
100  lb.  of  60-per-cent  gelatin  powder.  The  side  and  back 
holes  were  then  fired,  using  from  50  to  70  lb.  of  40-per  cent 
gelatine  powder.  In  no  case  were  the  holes  tamped. 

The  greatest  footage  made  in  any  one  month  under  this 
system  and  management  was  267.6  ft.,  and  a  total  of  2760  was 
made  from  September,  1899,  to  August,  1900,  at  an  average 
cost  of  $28  per  foot,  or  a  total  of  $79,470. 

The  premium  system  of  wages  for  employes  was  used.  This 
consisted  of  paying  $6  for  every  foot  driven  over  160  ft.  per 
month.  This  $6  was  divided  proportionately  between  the  drill 
gang,  powder  gang,  and  muckers,  according  to  their  wages. 

The  success  of  the  rapid  progress  of  the  tunnel  under  this 
management  can  be  attributed  to  the  following  points: 

1.  The  use  of  high-pressure  air,  160  lb.,  being  the  minimum 
pressure  kept  at  the  breast. 

2.  The  use  of  the  Water  Leyner  machine  for  the  back  holes ; 
for  by  putting  in  these  holes  to  a  greater  depth,  it  was  possible 
to  put  in  a  much  deeper  round  throughout. 

3.  The  premium  system  of  paying,  which  offered  an  incen- 
tive to  employes  to  do  their  best  work. 


MISCELLANEOUS  INSIDE  COSTS  421 

Below  is  given  a  typical  monthly  footage  expense  during 
the  year  above  mentioned  taken  from  the  1900  annual  report : 

Drill-crew  men  and  foreman $3.11 

Trammers  and  drillers :.  . .  .     3 . 88 

Blacksmith  shop 1 . 13 

Engineers 1 . 08 

—    $9.20 

Ammunition 4 . 09 

Oil  and  waste .18 

Coal  for  power .  3 . 78 

Feed  and  shoes  for  mules .23 

Drill  repairs .81 

Premiums  on  footage .61 

Track  equipment  and  repairs .78 

Mine  timber .31 

Labor 1.92 

Material 1 . 10 

Labor,  repairs  along  tunnel 1 . 29 

Surveying .21 

Legal  expense '. .19 

Insurance .07 

Salary  and  office  expense 2 . 39 

Minor  expenses .58 


Total $27.74 

Some  interesting  cost  data  were  obtained  in  driving  a  rock 
tunnel  at  the  property  of  the  Iron  Mountain  Tunnel  Co.  at 
Superior,  Mont.,  in  1906  and  1907. 

The  new  tunnel  is  7  X  6  ft.,  on  a  grade  of  4  in.  per  100  ft., 
and  has  a  drainage  flume  12  X  12  in.  laid  in  the  flooring.  The 
work  of  driving  the  tunnel  which  is  to  be  5602  ft.  long  was 
begun  February  10,  1906.  In  October,  1906,  it  had  been  driven 
1500  ft.,  and  4345  ft.  had  been  excavated  by  September,  1907, 
leaving  uncompleted  a  distance  of  1255  ft. 

Excepting  four  stretches  of  ground,  aggregating  in  all  about 
800  ft.,  the  rock  encountered  was  very  hard,  requiring  a  heavy 
expenditure  of  explosives.  In  the  distance  the  tunnel  has  thus 
far  been  driven,  it  has  been  necessary  to  timber  only  535  ft., 
the  formation  requiring  support  being  encountered  in  four 
different  places,  one  of  them  325  ft.  long. 

The  accompanying  table  gives  the  monthly  costs  of  the 
work  up  to  September,  1907 : 


422 


COAL  MINING  COSTS 


Time  of 
Driving 

Feet 

COST  PER  FOOT 

FEET  DRIVEN 

Gross 

Driving 
and 
Equipping 

Driving 

Side- 
track 

Cross- 
cuts 

Prior  to  June, 
1906   

559 

231 

215 
229 
227 

285 
264 
260 

288 
267 
244 
251 
254 
251 
206 
228 
170 

$20.11 
14.81 
14.96 
13.29 
12.95 
12.66 
15.77 

12.08 
12.09 
13.81 
13.03 
14.35 
13.37 
14.65 
18.11 
15.95 

$18.70 
11.81 
14.85 
12.31 
12.59 
12.30 
15.36 

11.79 
11.76 
13.44 
12.68 
13.99 
13.00 
14.22 
17.67 
15.55 

$15.58 
11.01 
14.15 
11.59 
11.80 
11.00 
13.39 

10.64 
10.69 
12.75 
11.53 
13.03 
11.88 
13.02 
16.27 
14.45 

40 
80 

85 

100 
60 
50 
70 
43 
50 
64 

100 

10 

9 

4 

1906 
June               .  .  . 

July 

August  
September  
October     

November  
December  

1907 
January  

February  

March  
April  

May  
June  

July 

August  
September.  . 

Totals  
Averages  

4429 
233.1 

$231.99 
14.50 

$222.02 
13.88 

$202.78 
12.67 

742 
46§ 

23 
H 

NOTE. — No  report  was  made  of  costs  per  foot  of  the  first  three  months' 
operation. 

The  city  of  Los  Angeles,  in  California,  started  constructing 
an  aqueduct  about  217  miles  long  in  1909,  including  105  tunnels 
whose  aggregate  length  is  28  miles.  The  Elizabeth  tunnel, 
which  is  26,860  ft.  long,  was  driven  from  both  ends,  termed 
the  north  and  south  portals.  This  tunnel  has  a  cross-section 
12  ft.  4  in.  X  12  ft.  9  in.,  a  grade  of  1  ft.  in  1000  ft.,  and  a 
water  capacity  of  1000  sec.-ft 

The  general  equipment  at  this  tunnel  consists  of  the  follow- 
ing: 

Two  compressors,  520  ft.  capacity,  belt  driven  from  elec- 


MISCELLANEOUS  INSIDE  COSTS  423 

trie  motors;  two  motor-generator  sets,  150  hp. ;  one  50-hp. 
electric  locomotive;  one  30-hp.  electric  locomotive;  nine  water 
Leyner  drills ;  38  rocker  dump  cars,  32  cu.f t.  capacity ;  one  drill 
sharpener. 

In  addition,  the  machine-shop  equipment  included  a  lathe, 
drill  press,  saws,  blowers,  motors,  and  the  necessary  tools  for 
such  work.  Most  of  the  destructible  equipment  was  supplied 
in  duplicate,  and  extra  machine  drills  were  furnished.  Each 
shift  was  supplied  with  a  tool  box  and  all  the  tools  necessary 
for  its  members'  work.  These  tool  boxes  were  locked  and  one 
man  on  each  shift  was  held  responsible.  A  station  was  cut  in 
the  tunnel  where  all  repairs  to  machines,  hose,  tools,  etc.,  were 
made.  Wherever  possible  each  individual  was  held  responsible 
for  the  tools  he  used. 

Each  shift  was  required  to  drill  and  blast,  the  length  of 
round  being  regulated  by  the  nature  of  the  ground  encoun- 
tered in  the  first  hole  drilled.  Discipline  and  strict  attention 
to  business  while  on  shift  were  required  of  everyone,  while  at 
the  same  time  a  spirit  of  friendliness  was  fostered  and  every 
man  made  to  feel  that  in  a  large  measure  the  success  of  the 
project  was  due  to  his  own  efforts  and  the  interest  he  took  in 
the  work. 

A  bonus  of  40c.  per  foot  per  man  for  every  foot  driven 
over  22/3  ft.  per  shift  was  paid,  bringing  the  wages  up  to  a 
good  figure  above  the  scales  usually  paid  in  other  mining 
camps  and  consequently  bringing  a  better  grade  .of  men  than 
could  have  otherwise  been  obtained. 

The  rock  encountered  was  mainly  granite,  which  in  places 
shades  into  both  gneiss  and  schist.  The  granite  is  composed 
mainly  of  feldspars  and  biotite  with  only  a  small  amount  of 
quartz  usually  present.  Its  texture  varies  at  different  points 
along  the  tunnel,  the  finely  crystalline  rock  usually  being  with- 
out joints  or  seams  and  the  coarsely  crystalline  rock  usually 
being  full  of  seams.  The  finer  crystalline  rock,  therefore, 
allows  much  the  slower  progress  in  tunnel  driving,  while  the 
blocky  ground  gives  about  as  near  ideal  conditions  for  record- 
breaking  drives  as  one  is  apt  to  find. 

The  rock  in  the  south  heading  is  hard,  unaltered,  and  in 
general  blocky  enough  to  afford  fine  breaking  qualities.  It 
requires  no  timbering  except  one  or  two  sets  at  long  intervals 


424 


COAL  MINING  COSTS 


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426  COAL  MINING  COSTS 

where  the  roof  is  heavy,  and  there  is  comparatively  little  water 
encountered. 

In  the  north  heading,  the  rock  has  been  shattered  by  scores 
of  fissures,  and  in  general  is  so  kaolinized  by  percolating  waters 
that  it  is  for  the  most  part  soft,  friable  and  frequently  pasty. 
Considerable  water  enters  the  tunnel,  and  the  ground  is  so 
heavy  it  requires  timbering,  except  at  one  or  two  places  in 
the  tunnel  of  small  extent  where  hard  unaltered  rock  was 
penetrated.  Long  extents  of  heavy,  treacherous  ground  are 
continually  encountered,  so  that  the  timbering  must  be  kept 
well  up  to  the  face. 

As  a  result  of  these  widely  varying  conditions  neither  the 
costs,  nor  the  record  performances  of  work  at  one  heading 
may  be  fairly  taken  as  a  criterion  of  the  work  at  the  other 
heading  of  this  same  tunnel. 

In  spite  of  this  diversity  of  work,  however,  the  progress 
made  at  each  heading  was  surprisingly  uniform  as  shown  in  the 
accompanying  table,  until  at  the  end  of  April,  1910,  30  months 
after  starting  work,  the  north  heading  had  penetrated  9754  ft. 
and  the  south  heading  9738  ft.,  totaling  19,492  ft.,  or  nearly 
72  per  cent  of  the  entire  length. 

To  those  interested  in  the  rivalry  between  the  crews  of  the 
north  and  south  Elizabeth  tunnel  headings,  the  accompanying 
tables  will  be  replete  with  interest  as  well  as  with  valuable 
data.  These  give  the  official  costs  for  both  headings  during 
the  month  of  April,  1910,  when  the  south  portal  advanced  the 
American  record  to  604  ft.,  while  the  north  portal  progressed 
561  ft.  in  spite  of  exceptionally  heavy  ground.  It  will  be 
noted  that,  as  might  be  expected,  the  cost  of  explosives  was 
the  heavier  for  the  south  heading,  while  the  cost  of  timbering- 
is  mainly  responsible  for  increasing  the  cost  per  foot  of  the 
north  heading  to  $4.36  more  than  the  $25.25  of  the  south  head- 
ing. It  is  also  significant  that  the  cost  of  mucking  in  the  two 
headings  is  very  nearly  the  same. 

The  appended  study  covers  unit  costs  of  driving  a  timbered 
tunnel  (North  Portal,  Tunnel  No.  7)  in  the  little  Lake  sub- 
division of  the  Grapevine  division  of  the  Los  Angeles  aque- 
duct. 

Ninety  feet  were  driven  in  15  eight-hour  shifts,  the  period 
covered  by  detailed  cost  keeping. 


UNIT  COSTS 

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428 


COAL  MINING  COSTS 


SOUTH  PORTAL  ELIZABETH  TUNNEL.     DETAILS  OP  ANNUAL  SUMMARY  OF 
TUNNEL  REPORTS  FOR  1909 


Totals 

Units 

Total 
Cost 

Unit 
Cost 

Footage  to  date 

7,585  ft. 

Footage  during  1909  
Daily  footage  (365  days)  . 

4,476  ft. 

12  26 

Footage  of  untimbered  section  
Actual  cost  of  untimbered  section  .  .  . 

3,274  ft. 

121  787  45 

37  198 

Footage  of  timbered  section  
Actual  cost  of  timbered  section  

1,202  ft. 

51  941  11 

43  20 

Average   cost   of   timbering  per  foot 
progress               

1  16 

Total  bonus  footage 

1,563  ft. 

Total  bonus  pay  roll  .    . 

23440  23 

Average  cost  of  bonus  per  foot  progress 
Actual  expenditure  



173  728  56 

.523 

Number  of  shifts  worked  
Number  of  shifts  lost  
Number  of  men  days.  .  . 

1,055 
40 
21,452 

Men  per  foot  progress  
Average  progress  per  shift 

4.79 
4  24 

Number  of  holes  drilled 

21  066 

Total  feet  drilled     .  . 

135,896 

Feet  drilled  per  foot  progress 

30  36 

Total  time  drilling  (in  hours)  

3,656  25 

Average  time  drilling  per  foot  progress 
(in  minutes)  

49  01 

Average  time  drilling  one  hole  1  foot 
deep 

1  61 

Pounds    of    powder    used    (including 
trimming)  

143,659 

Average  pounds  per  foot.  .  . 

32  09 

Number  of  cars  mucked  (32  cubic  feet). 
Cars  mucked  per  foot  

28,561 

6  40 

Leyner  No.  9,  drill  repairs  (3  machines 
per  shift)   

4  116  99 

Average  cost  of  repairs  per  machine 
per  foot  .          

30 

Leyner  No.  2  drill  sharpener  repairs 
(for  3300  feet) 

387  13 

117 

Energy  used  (kilowatt  hours)  

608,762 

135.00 

MISCELLANEOUS  INSIDE  COSTS 


429 


SOUTH   PORTAL   ELIZABETH  TUNNEL.     SUMMARY  OF  TOTAL  EXPENDITURE 

DURING  1909 


Totals 

Unit  Cost 

A.  Engineer  and  superintendent  
B.  Excavation  
C    ]M  uc  king 

$3,851  .  24 
79,607.15 
55,556  .  33 

$0.86 
17.78 
12.41 

E    Drainage           

421.01 

0.09 

F    Ventilation                          

2,313.44 

0.52 

G   Light  and  power 

24,750.71 

5.53 

Total  cost  untimbered  

$166,499  .  88 

$37.19 

D   Timbering 

7,228  68 

6.01 

Total  cost  of  timbered  tunnel 

$173,728.56 

$43.20 

SOUTH  PORTAL  ELIZABETH  TUNNEL.     DETAILS  OF  TUNNEL  REPORTS, 
APRIL,  1910 


Totals 

Units 

Total 
Cost 

Unit 
Cost 

Total  footage  to  date  
Footage  during  April,  1910  

9,738  ft. 
604ft. 

Daily  footage  (30  days) 

20  13 

604  ft. 

Actual  cost  of  untimbered  section.  .  . 

15,257.61 

25.25 

Footage  of  timbered  section  

000  ft. 

Bonus  footage    

364  ft. 

Bonus  pay  roll                           .    ... 

2,377.48 

3.93 

78,379 

129.76 

Cost  of  energy,  at  1.85  c.  per  kilo- 
watt hour  .  .        

1,450.02 

2.40 

*Estimated  expenditure  

24,238.52 

Actual  expenditure    . 

15,257.61 

*Amount  saved  over  estimated  ex- 
penditure 

8,980.91 

Number  of  underground  men  days: 
Foreman  and  heading  crew  
Timber,    pipe,    track,    car,    and 
machine  repair  men 

1,860 

458 

3.079 

.758 

Mechanics,  electrician  and  helpers. 

154 

.255 

430 


COAL  MINING  COSTS 


SOUTH  PORTAL  ELIZABETH  TUNNEL.     DETAILS  OF  TUNNEL  REPORTS, 
APRIL,  1910 — Continued 


Totals 

Units 

Total 
Cost 

Unit 
Cost 

Number  of  mule  davs 

90 

81.00 

0.134 

Number  of  shifts  worked  

90 

Average  progress  per  shift  
Number  of  holes  drilled 

1,924 

6.711 

Total  feet  drilled                 

16,079 

Feet  drilled  foot  progress 

26.62 

Total  time  drilling  heading  (hours).  . 
Average  time  drilling  heading  per 
foot  (minutes)  ...        

324 

32.18 

Average  time  drilling  one  hole  one 
foot  (minutes) 

1.209 

Pounds  of  powder  used   (including 
trimming)         

16,100 

26.65 

Number  of  cars  mucked  (32  cu.  ft.) 
heading  and  ditch 

3,216 

5.324 

Drill  repairs  (three  No  9  Leyners) 

248  .  77 

Average    cost   of   repairs   per   foot 
(three  machines) 

0.411 

Average    cost   of    repairs   per   foot 
(one  machine) 

0.137 

Drill     sharpener     repairs     (Leyner 
No.  2)  

61.70 

0.102 

Drill  steel  broken                        

380 

.629 

Drill  steel  sharpened  
Drill  steel  welded              

6,865 
445 

11.36 

.736 

Cars  repaired                                

38 

.062 

Cars  repairs  

48.52 

0.8 

Car     equipment     (changing     from 
12-inch  to  14-inch  wheels) 

260  68 

0  431 

NOTE.— New  American  hard  rock  tunnel  record  established  April,  1910,  604  ft. 


MISCELLANEOUS  INSIDE  COSTS 
SUMMARY  OF  GENERAL  EXPENDITURES  DURING  1909 


431 


Total  Cost 

Miscellaneous  structures  

$165  79 

Tunel  construction 

15  348  27 

Miscellaneous  construction  equipment  

5,168.85 

Miscellaneous  camp  eouipment 

19  22 

M.  &  O.  live  stock  

183  75 

M.  &  O.  local  telephone  lines 

31  73 

M.  &  O.  water  supply. 

2  34 

Division  administration  ... 

740  86 

M.  &  O.  roads  and  trails. 

116  05 

Total  expenditure  

$21,776  86 

Pay  roll  

9  419  76 

Bonus  roll 

2  377  48 

Material  issues  

6  915  83 

Material  receipts.  . 

2.841.54 

SUMMARY  OF  TUNNEL  EXPENDITURE  DURING  APRIL,  1910 


Totals 

Units 

E.  W.  O.  33  —  A.  Engineering  and  superintendent.  .  .  . 
B.  Excavation. 

$162.89 
7  163  30 

.27 
11  86 

C.  Mucking  
E.  Drainage.  .  .  . 

5,275.78 
44  49 

8.73 
07 

F.  Ventilation 

85  10 

14 

G.  Light  and  power  

2,109  55 

3  49 

D.  Timbering.  .  . 

14,841.11 
90  66 

24.56 

K.  Back  trimming. 

416  50 

69 

Total  expenditure 

$15348  27 

25  25 

The  tunnel  is  approximately  10  X  10  ft.  in  section ; 
cu.yd.  in  place  per  linear  foot  to  pay  line ;  overbreakage  about 
17  per  cent,  making  a  total  of  61/2  cu.yd.  of  broken  material 
per  foot  of  tunnel. 

The  heading  was  in  800  ft.,  lighted  by  electricity  at  110 
volts,  ventilated  by  a  No.  3  Champion  blower  through  12-in. 
pipe,  the  heading  being  cleared  in  15  min.  after  shooting. 


432  COAL  MINING  COSTS 

Drilling  is  done  by  one  No.  7  Leyner  drill,  water  being 
forced  through  hollow  steel;  drill  uses  approximately  66  cu.ft. 
free  air  per  minute  at  83  Ib.  pressure  per  square  inch,  drilling 
holes  to  10  ft.  in  depth. 

Mucking  is  accomplished  by  use  of  steel  sheets  laid  down 
before  shooting;  No.  3  D-handle,  square-point,  shovels,  and 
32-cu.ft.  rocker  dump  cars  pulled  by  a  3y2-ton  locomotive, 
running  on  a  24-in.  gauge  single  track  laid  with  25-lb.  steel. 

The  rock  is  a  close-grained,  hard,  gray  granite  with  numer- 
ous seams,  causing  the  drill  to  run  from  alignment,  but  breaks 
well.  The  seams  and  water  combined  make  it  necessary  to 
timber  all  this  ground.  The  ground  carries  enough  water  to 
make  disagreeable  mucking,  and  has  to  be  pumped  out. 

Timbers  are  of  6  X  8  in.  Oregon  pine  spaced  5  to  8  ft.  apart, 
as  ground  permits,  lagged  with  2  X  6  in.  plank.  Sets  of  timbers 
consist  of  two  vertical  posts  and  a  four-segment  arch. 

The  crew  consisted  of  1  shift  boss  at  $3.50  per  day;  4 
miners  at  $3.50 ;  5  muckers  at  $2.50,  and  1  trammer  at  $2.50. 
The  blacksmith  doing  repair  work  was  paid  $4  per  day. 

The  four  miners  worked  on  day  shift  drilling  the  ground, 
timbering  and  shooting,  the  muckers  following  on  night  shift, 
resulting  in  a  clean  heading  for  the  drill  crews,  and  nothing 
interfering  with  the  mucking  crew. 

In  February,  1908,  the  Chamber  of  Mines  and  the  Trans- 
vaal Government  jointly  offered  a  prize  amounting  to  £7500 
for  the  best  drilling  machine  that  could  be  produced  for  mine 
work  on  the  Witwatersrand.  There  were  23  entries  for  this 
prize,  19  of  which  started  in  the  competition. 

In  the  elimination  trials,  which  were  carried  out  on  the  sur- 
face at  the  Transvaal  University  College  and  underground  at 
the  Ferreira  Deep,  Ltd.,  nine  drills  were  eliminated  and  10 
entered  for  the  competition  of  300  shifts.  These  were  the  Hoi- 
man  2%  in.,  Holman  2%  in.,  Siskol,  Climax  Imperial,  New  Cen- 
tury 00,  Konomax,  Chersen,  Waugh,  Murphy,  Westfalia. 

In  the  surface  elimination  test  the  most  rapid  rate  of  drill- 
ing was  accomplished  by  the  Westfalia  machines;  namely, 
4.996  in.  per  minute  of  actual  drilling  time.  The  largest  air 
consumption  recorded  was  that  of  one  of  the  Konomax  ma- 
chines; namely,  125.9  cu.ft.  of  free  air  per  foot  drilled. 


MISCELLANEOUS  INSIDE  COSTS 
LABOR  COSTS 


433 


Class  of  Work 

Total 
Hours 
Labor 

Total 
Labor 

Costs 

Cost  per 
Foot  of 
Tunnel 

Inside  Labor: 
Squaring  heading.  
Setting  up  and  tearing  down  machine  . 
Drilling,  one  No.  7  Leyner;   including 
shift  boss'  time  

23.50 
36.00 

55.33 

$  9.03 
16.59 

43.21 

$0.100 
0.184 

0.480 

Number  of  holes  drilled  150 

Total  footage  of  holes  1,202.30 
Feet  drilled  per  hour,  includ- 
ing lost  time  15  .  84 

Feet  drilled  per  hour,  actual 
drilling  time  21  .  74 

Average  depth  of  holes,  feet  .  .        8  .  00 
Cost  per  foot  of  hole,  cents  ...        3  .  60 
Fastest  hole  9  ft.  6  in.  in  10  min  
Slowest  hole  8  ft.  6  in.  in  1  hour  18  min. 
Average  hole  8  ft  in  22  min    

Blowing  out  holes                     

5.75 

4.15 

0.046 

Loading  and  shooting                        .  .  . 

56.25 

22.31 

0.248 

Mucking  412  cars                                .  . 

835  .  00 

268.44 

2.980 

Trimming,  stulling,  caves,  etc  
Timbering  (cost  per  M.  ft.  =$11.32)..  . 
Lost  time                                       

102.00 
107.25 
40.75 

39.24 
40.82 
15.66 

0.436 
0.453 
0.174 

Bonus,  30  c.  per  man  per  foot  in  excess 
of  2  3  ft  per  shift                    

112.08 

1.240 

Repairs,  to  trolley,  pump,  etc  

3.25 

1.20 

0.013 

Totals  inside  labor 

1,265  08 

572  73 

6.354 

Outside  Labor: 
Sharpening  steel,  with  Leyner  No.  2 
machine 

44  00 

17  83 

0.198 

Repairing  drill 

7  50 

2  88 

0.032 

Framing  timbers,  at  shop,  per 
M  ft                                           =2  42 

8  75 

0.097 

Light  and  power  

90.00 

33.75 

0.375 

Totals  outside  labor 

141  50 

63  21 

0  702 

Auxiliary  Labor: 
Laying  track  90  ft 

31  50 

12  75 

0  141 

Drainage  line  90  ft 

43  50 

14  33 

0  160 

Ventilation  line  72  ft 

11  00 

3  44 

0.048 

Trolley  line  95  ft 

18  00 

6  35 

0.067 

Air  line  80  ft                                      ... 

2  50 

0.80 

0.010 

Water  line  80  ft                               .... 

2  50 

0  79 

0.010 

Lights  line,  90ft  

8.00 

2.86 

0.032 

Totals  auxiliary  labor 

117  00 

41  32 

0.468 

Local  Administration  and  Engineering: 
Proportion   of   division   engineer   and 
assistant's  time  

50.40 

0.560 

Total  labor  costs 

1,523  58 

$727  66 

$8  .  094 

434 


COAL  MINING  COSTS 


COST  OF  MATERIALS  AND  SUPPLIES 


Total 

Material 

Costs 


Construction  Materials  and  Supplies: 

Drill  repairs,  2  side  rods,  $1.40;    1  chuck,  $15; 

2  rings,  $2.02;   1  oil  can,  .15;  1  belt,  .16;  20  per 

cent,  freight,  $3.74 $  22.47 

Cost  per  foot  of  hole  =  .018 
Drill  supplies,  machine  oil,  .58;  drill  steel,  45  ins., 

$4.12;    412  Ib.  blacksmith  coal,  $3.14;    20  per 

cent  freight,  $1.57 9.41 

Cost  per  foot  of  hole  =  .008. 
Mucking  supplies,  car  oil,  .20;    pick  handle,  .26; 

hammer  handle,  .15;  20  per  cent  freight,  .12. ...  .73 

Power,  machine,  2052  kw.  h.;   blower  and  lights, 

355  kw.  h.;   locomotive,  1800  kw.  h.=4207  kw. 

hours,  at  1.7  c 71 . 52 

Explosives,    tamping   stick,    .40;    2700   ft.   fuse, 

$11.54;    306  15  gr.  Lion   caps,  $2.08;    650  Ib. 

1|    inch,    40-per-cent   gelatin   powder,    $69.88; 

250  Ib.  1  in.  40-per-cent  gelatin  powder,  $26.88; 

150  Ib.  1  in.,  60-per-cent  gelatin  powder,  $20.63; 

20  per  cent  freight  ,$26.28  157.69 

1050  Ib.  powder  =  11. 66  Ib.  per  foot  of  tunnel. 
1050  Ib.  powder  =  3. 3  Ib.  per  cubic  yard  in  place. 
Explosive  cost  =  $.050  per  cubic  yard  in  place. 
Explosive  cost  =  $.27  per  cubic  yard  broken. 
Timbers,  3597  feet  B.  M.  lumber,  $59.24;  freight 

on  same  $55.01;    775  wedges,  $5.43;   50  dowel 

pins,  .42;  nails,  .73;  freight,  $1.33 122. 16 

Timber  per  foot  of  tunnel,  B.  M.=40. 
Lighting,  candles,  $7.15;    14,  16,  and  32  candle- 
power  globes,  $2.58;  20  per  cent  freight,  $1.94.  .        11.67 

Totals  for  construction  materials 395 . 65 

Auxiliary  Material: 
Trackage,   180  feet,  25  Ib.  rail,  $15;  splices  and 

bolts,  .83;   spikes,  .18;  ties,  $1.42;  20  per  cent 

freight,  $3.68 21 . 11 

Drainage  90  ft.,  250  ft.,  wire,  $2.13;   knobs,  .07; 

2-inch  pipe,  $4.36;  20  per  cent  freight,  $1.31 7.87 

Ventilation  72  ft.,   12-inch  pipe,  $24.34;    20  per 

cent  freight,  $4.87 29.21 

Trolley  95  ft.,  wire,  $6.32;    lumber,   .37;    fittings, 

$1.42;  20  per  cent  freight,  $1.62 9 . 73 

Air  line  80  ft.,  pipe,  $5.28;    fittings,  .76;    20  per 

cent  freight,  $1.21 7.25 

Water  line,  80  ft.,  pipe,  $5.28;  fittings,  .76;  20  per 

cent  freight,  $1.21 7 . 25 

Light  line,  90  ft.,  wire,  $1.48;  fittings,  .37;  20  per 

cent  freight,  37 _ 

Total  auxiliary  materials 84 . 64 

Total  material  costs 480 . 29 

Live  Stock: 
Mule  15  days  at  90  c $13.50 

Total  direct  and  auxiliary  field 

charges $1,523.58     $727.61    $480.29 


MISCELLANEOUS  INSIDE  COSTS 
RECAPITULATION.     COSTS  PER  FOOT  OF  TUNNEL 


435 


Labor,  direct  charge $  7 . 056 

Material  and  supplies,  direct  charge 4 . 399 

Local  administration  and  engineering 0 . 560 

Stock  service 0. 150 

12.165 

Labor  on  tracks,  etc 468 

Material  for  tracks,  etc 1 . 034 

1.502 
As  this  work  will  salvage  at  about  66  per  cent,  we  deduct . .    .  690 

Net  charge  for  auxiliary  work 0 . 812 

Estimated  proportion  of  charge  for  roads  and  trails  on  division . .  1 . 500 

Estimated  proportion  of  charge  for  buildings  on  division 0 . 200 

Estimated  proportion  of  charge  for  water  supply  on  division ....  0 . 220 

Estimated  proportion  of  charge  for  machinery  and  tools 1 . 060 

Total  field  charges 15.957 

Add  3  per  cent  to  cost  for  executive  office  administration 0 . 475 

Total  cost  of  tunnel  ready  for  lining $16 . 432 


In  the  underground  elimination  air  trials  the  quickest  drill- 
ing speeds  were  attained  by  the  Chersen  and  Holman  2%-in. 
machines.  The  former  drilled  1.81  in.  per  minute  over  drilling 
and  changing  time  and  1.56  in.  per  minute  over  total  time, 
which  consisted  of  three  8-hr,  shifts.  The  figures  recorded 
for  the  Holman  2%-in.  machine  on  this  trial  were  1.94  and 
1.47  in.  per  minute,  respectively. 

As  the  competition  proceeded  the  following  machines  with- 
drew :  Climax  Imperial,  Konomax,  Murphy,  Waugh,  and  West- 
falia.  It  was  the  intention  to  run  300  shifts,  but  owing  to 
lack  of  air  pressure,  215  shifts  were  run.  The  stopes  drilled 
in  were  24  to  45  in.  wide.  The  number  of  feet  drilled  by  each 
pair  of  the  four  leading  competing  machines  were  as  follows: 
Holman  2y8-in.,  12,779  ft. ;  Siskol,  14,083  ft. ;  Holman  2%-in., 
11,744  ft. ;  Chersen,  11,781  ft.  This  was  drilled  in  215  shifts 
of  8  hr.  each. 

The  drilling  speeds  attained  over  total  times  by  the  four 
machines  are  as  follows : 


436 


COAL  MINING  COSTS 


Holman,  2£  inches 742  in.  per  min. 

Siskol 818  in.  per  min. 

Holman,  2f  inches 682  in.  per  min. 

Chersen 684  in.  per  min. 

At  the  surface  elimination  trials  the  average  rate  of  drilling 
of  each  pair  of  machines  was  as  follows : 


Machine 

Total  Time 
Inches  Per  Minute 

Actual  Drilling  Time 
Inches  Per  Minute 

Holman  2|  in  
Siskol 

1.566 

2  058 

2.393 
4.337 

Holman  2f  in  
Chersen 

1.988 
2.515 

3.110 
4.110 

The  cost  of  drilling  per  foot  with  these  four  leading  ma- 
chines amounted  to:  Holman  2%-in.,  9.77d. ;  Siskol,  9.90d. ; 
Holman  2%-m.,  10.91d.;  Chersen,  11.94d.  These  figures  com- 
pare favorably  with  the  cost  of  hand  drilling  on  the  Rand. 

In  the  competition,  two  sets  of  machines,  the  Holman 
2%-in.  and  the  Siskol  have  cost  approximately  8.3d.  per  foot 
drilled  plus  the  cost  of  steel  and  drill  sharpening,  which  would 
come  to,  at  the  most,  1.5d.  per  foot,  making  9.8d.  per  foot. 
But  it  has  taken  1.282  mine  shifts  to  make  one  8-hr,  shift, 
therefore  the  wage  cost  must  be  increased  in  that  ratio.  Fur- 
ther, instead  of  assuming  10s.  per  shift  for  two  machines  as 
the  white  wage,  25s.  per  shift  were  taken  for  four  machines. 
This  alteration  means  an  increase  of  y2d.  per  foot  drilled,  mak- 
ing the  total  cost  per  foot  of  these  two  sets  of  machines  prac- 
tically 11. 8d.  per  foot.  This  is  as  near  as  one  can  get  to  the 
actual  cost  per  foot  for  28,528  ft.  drilled  by  these  machines. 
This  cost  of  11. 8d.  per  foot  compares  favorably  with  the  cost 
per  foot  by  native  labor,  for  which  Is.  Id.  per  foot  would  be  an 
exceedingly  low  figure.  The  tonnage  broken  would  be  as  6  to 
5  in  favor  of  the  machines.  The  cost  of  explosives  would  be 
as  6  to  5  in  favor  of  native  labor.  The  tendency  is  for  native 
wages  and  cost  to  increase  and  this  is  the  heavy  item,  lOd.  per 
foot,  in  hammer  work,  but  a  much  smaller  item,  only  3d.  per 
foot  in  machinery,  With  a  proper  size  of  bit  a  hole  4  ft.  8  in. 


MISCELLANEOUS  INSIDE  COSTS 


437 


or  5  ft.  deep  is  much  better  as  regards  tonnage  than  the  ordi- 
nary hand-drill  hole,  for  it  will  carry  more  explosives  and  so  a 
greater  burden. 


Inches  Drilled 

Per  Pound. 
Weight  of  Machine 
Over  Drilling 

Depth  of 
Holes  Drilled 

Percent- 

Period of  4  Hours 

age  In- 

Weight 

crease  in 

Name  of  Drill 

of  Drill 
in 
Pounds 

Depth 
Using  60 
Pounds 

Air  Pres- 
sure 50 

Air  Pres- 
sure 60 

Air  Pres- 
sure 50 

Air  Pres- 
sure 60 

Pounds 

Pounds 

Pounds 

Pounds 

Air 

Per 

Per 

Per 

Per 

Pressure 

Square 

Square 

Square 

Square 

Inch 

Inch 

Inch 

Inch 

Ft.      In. 

Ft.      In. 

Flottman  

52.250 

4.33 

5.83 

18     lOf 

25     4f 

34.4 

Gordon 

72  .  625 

4.69 

6.08 

25      41 

36    9£ 

29.5 

Holrnan 

97.500 

1.77 

2.09 

*****            ^*-$ 

14      4J 

17    Oi 

18.6 

Kimber  

100.000 

1.55 

2.03 

•*-4: 

12     llf 

W4 

16  11 

30.6 

Little  Kid  

102.500 

1.85 

2.63 

15      9i 

22     6} 

42.6 

Chersen 

113.625 

2.79 

3.50 

26      5f 

33     2 

26.0 

Little  Wonder.... 

118.500 

1.68 

2.02 

ArfVF                   V  4 

16      71 

19  llf 

19.7 

Baby  Ingersoll.  .  . 

129.500 

2.34 

2.74 

25      3 

29    6} 

17.0 

Drill  steel. — There  is  perhaps  no  single  item  in  rock  drill- 
ing that  will  affect  the  cost  of  the  work  so  quickly  as  the  effi- 
ciency of  the  drill  steels.  The  essential  qualities  of  a  drill 
steel  as  laid  down  in  a  paper  presented  at  the  February,  1921, 
meeting  of  the  Am.  Inst.  of  Min.  and  Met.  Engr's.,  are:  First, 
it  must  be  easily  forged;  second,  the  forged  bit  end  must  be 
such  that  it  can  be  easily  heat  treated  to  obtain  hardness  to 
resist  chipping;  third,  the  bar  or  body  must  be  stiff  to  resist 
bending  or  twisting  and  yet  tough  to  resist  shock  and  vibra- 
tion, with  resulting  breakage,  and  fourth,  the  forged  shank  end 
must  be  such  that  it  can  be  easily  heat  treated  to  obtain  some 
hardness  with  great  toughness. 

Drill  steel  must  be  properly  forged  either  by  hand  or 
machine,  and  this  operation  requires  pyrometric  control.  Two 


438  COAL  MINING  COSTS 

causes  of  trouble  can  be  eliminated  at  this  point,  as  there  is  no 
doubt  "that  drill  steels  as  a  rule  are  forged  at  too  high  tem- 
peratures, and  the  forging  operation  is  continued  after  the 
temperature  has  dropped  below  the  critical.  Furthermore, 
little  if  any  annealing  is  done  on  drill  steel  after  forging; 
hence  the  forging  operation  must  be  conducted  with  great 
care. 

Assuming  the  forging  temperature  is  correct,  the  other 
minor  requirements  are  that  the  bit  and  the  shank  be  in  align- 
ment with  the  body ;  that  the  shank  shall  be  of  the  proper  shape 
and  length  and  the  shank  collar  or  lugs  be  of  the  proper 
diameter  and  length ;  that  the  hole,  if  any,  be  free  from  obstruc- 
tion; that  the  striking  end  of  the  shank  be  flat  and  square; 
that  the  bit  be  of  the  proper  shape,  with  the  cutting  and  ream- 
ing edges  formed  full  and  to  the  required  size ;  that  the  ream- 
ing edges  are  concentric  with  the  axis  of  the  steel,  and  that 
there  are  no  sharp  corners  at  the  shoulder  where  the  bit  blends 
into  the  body. 

It  is  obvious  that  the  above  requirements  can  best  be 
obtained  day  in  and  day  out  by  means  of  a  drill  steel  sharpen- 
ing machine.  The  efficiency  of  such  a  machine  will  accordingly 
depend,  first,  on  the  initial  forging  temperature  required,  for 
the  lower  the  initial  forging  temperature  the  better  the  steel 
structure ;  second,  on  the  accuracy  of  the  forged  bit  and  shank ; 
third,  on  the  speed  of  operation ;  fourth,  on  restriction  of  the 
hole  in  the  shank  and  bit  when  using  hollow  drill  steel ;  fifth, 
on  the  number  of  heats  or  times  required  to  heat  the  steel  be- 
fore securing  the  finished  bit  or  shank,  and  sixth,  on  the  air 
consumption  of  power  required.  The  means  of  heating  for 
forging  will  be  considered  later. 

The  drill  steel  must  have  an  ideal  bit  and  shank.  The 
essential  qualities  of  such  a  bit,  regardless  of  the  conditions 
under  which  it  is  operated,  are  that  its  shape  be  such  that 
maximum  cutting  speed  can  be  maintained  for  as  great  a  dis- 
tance as  possible  before  wear  of  the  gauge  and  cutting  edge 
reduces  the  speed  of  penetration  to  a  point  where  a  change 
of  steel  is  made  necessary,  and  that  the  size  or  diameter  of  the 
drill  hole  corresponding  to  the  gauge  of  the  bit  can  be  main- 
tained with  the  least  possible  reduction  as  the  depth  of  the 


MISCELLANEOUS  INSIDE  COSTS  439 

hole  increases,  and  also  that  the  shape  of  the  bit  is  such  that 
it  can  be  correctly  and  readily  formed  and  heat  treated. 

The  following  features  of  bit  design  require  attention: 
Shape,  total  length,  and  angle  of  cutting  edge;  length  and 
area  of  the  reaming  edges  or  surfaces ;  size  and  shape  of  clear- 
ance grooves  for  ejection  of  cutting,  and  lengths  and  angle 
of  the  wings  and  the  manner  in  which  they  are  blended  into 
the  body  of  the  steel. 

The  combined  length  of  the  cutting  edge  and  the  manner 
in  which  it  is  applied  is  a  big  factor  in  the  drilling  operation 
regarding  the  speed  and  the  life  of  the  drill  steel  and  the  drill. 
The  longer  the  cutting  edge  the  greater  the  amount  of  rock 
cuttings  per  blow,  assuming  that  the  cutting  edge  is  and 
remains  sharp  or  sharp  enough  for  the  conditions.  Further- 
more, the  drilled  or  blunt  edge,  besides  decreasing  the  penetra- 
tion per  blow,  lessens  the  cushioning  effect,  and  this  causes  the 
drill  steel  to  rebound  from  the  rock,  which  may  cause  break- 
age of  the  drill  steel  or  parts  of  the  drill.  In  a  radial  cutting- 
edge  bit  the  work  done  is  greatest  at  the  extreme  cutting  and 
reaming  edges,  which  accounts  for  the  unequal  wear  along  the 
cutting  edge.  Therefore  it  is  apparent  that  the  only  way  to 
improve  this  condition  is  to  so  shape  the  cutting  edge  that 
the  work  is  evenly  distributed  throughout  its  length,  and  so 
that  the  extremities  of  the  cutting  edge  have  suitable  reaming 
surfaces  properly  tapered  back  to  improve  the  wearing  quali- 
ties of  the  gauge. 

A  good  deal  has  been  written  about  the  angle  and  shape 
of  the  cutting  edge,  and  many  kinds  of  bits  have  been  brought 
out,  such  as  the  bull  bit,  rose  bit,  double  cross  bit,  chisel  and 
double  chisel  bit,  and  other  types ;  but  the  bit  that  approaches 
the  ideal  design  is  the  double-arc,  double-taper  bit.  The 
accompanying  illustration,  Fig.  2  shows  the  characteristic 
dimensions  of  this  bit.  As  to  the  ideal  shank,  there  is  no 
doubt  that  the  so-called  shankless  steel  approaches  .the  ideal, 
and  it  is  regrettable  that  this  type  is  not  suitable  for  all  con- 
ditions. Its  advantages  and  disadvantages  are  evident. 

The  third  and  last  requirement  of  the  drill  steel  which  has 
ideal  qualities  is  that  it  must  be  properly  heat  treated.  This 
without  doubt  is  the  most  essential  operation  in  securing  best 
results,  and  yet  it  is  safe  to  say  it  receives  the  least  attention. 


440 


COAL  MINING  COSTS 


The  good  results  which  are  expected  from  all  previous  opera- 
tions can  be  obtained  only  through  proper  heat  treatment. 
Theoretically,  this  operation  of  heat  treating  should  be  simple; 
practically  it  is  not,  for  all  too  much  depends  on  the  equip- 
ment. 

It  is  interesting  to  note  that  answers  to  a  recent  questionnaire 
brought  out  the  following  facts  and  are  representative  of  all 
fields  and  conditions  of  mining. 

To  the  question,  what  type  of  bit  was  giving  the  best  service, 
returns  showed  that  the  double-arc,  double-taper  bit  and  cross 
bit  were  far  in  the  lead. 


This  end  rrtusfbe 
' 


FIG.  2. — Standard  design  for  double-arc,  double-taper  bit. 

To  the  question  as  to  what  necessitates  the  resharpening  of 
the  drill  steels  most  frequently,  wear  of  gage  was  unanimously 
first  choice ;  chipping  of  bit  was  unanimously,  with  one  excep- 
tion, second  choice;  breakage  of  shank  was  unanimously,  with 
one  exception,  third  choice;  upsetting  of  bit,  upsetting  of 
shank  and  breakage  of  body  were  about  on  a  par  for  fourth 
choice. 

To  the  question  as  to  what  was  the  direct  cause  of  the 
necessity  of  resharpening,  poor  heat  treatment  was  unani- 
mously first  choice,  and  faulty  steel  and  severe  rock  conditions 
were  about  on  a  par  for  second  choice. 

Explosives. — The  selection  of  explosives  for  tunnel  blasting, 
probably  requires  a  more  careful  study  of  conditions  than  for 


MISCELLANEOUS  INSIDE  COSTS  441 

any  other  kind  of  excavating.  Maximum  speed  and  economy 
in  driving  cannot  be  attained  unless  the  explosive  best  adapted 
to  the  work  is  used.  When  starting  a  tunnel  or  drift,  it  is  a 
good  plan  to  thoroughly  try  out  several  explosives,  which  are 
distinctly  different  in  action,  before  finally  adopting  any  one 
of  them.  The  results,  however,  from  this  preliminary  trial 
will  be  of  little  or  no  value,  unless  each  different  explosive  is 
used  under  exactly  the  same  conditions.  Care  must  be  taken 
to  see  that  no  change  occurs  in  the  character  of  the  rock, 
number  and  direction  of  the  bore  holes,  strength  of  the  de- 
tonator, kind  and  quantity  of  tamping,  amount  of  water 
encountered,  method  of  connecting  up  the  bore  holes  for  firing, 
and  that  the  explosive  is  always  thoroughly  thawed.  If  a 
material  change  in  any  of  these  conditions  occurs  as  the  work 
progresses,  further  tests  should  be  made  to  determine  whether 
a  quicker  or  slower,  a  stronger  or  weaker,  explosive  might  not 
break  the  ground,  or  bottom  the  bore  holes  better,  or  make  it 
possible  to  bring  out  the  cut  with  fewer  holes  or  deeper  ones. 
The  speed  at  which  rock  can  be  drilled  does  not  indicate  how 
it  will  break,  and  not  infrequently  that  which  can  be  easily 
drilled  is  very  difficult  to  blast. 

High  explosives  suitable  for  tunnel  blasting  should  not  give 
off  objectional  fumes  on  detonation,  and  accordingly  gelatin 
dynamite,  blasting  gelatin,  or  ammonia  dynamite  should  always 
be  selected. 

Gelatin  dynamite  is  made  in  various  grades  of  strength, 
from  25  to  80  per  cent,  inclusive.  It  is  comparatively  slow  in 
action,  the  higher  grades  being  little,  if  any,  quicker  than  the 
lower  ones. 

Blasting  gelatin  is  manufactured  in  only  one  strength, 
which  for  comparative  purposes  may  be  said  to  be  100  per 
cent.  It  is  more  powerful  and  quicker  acting  than  any  other 
blasting  explosive.  It  should  be  used  sparingly,  therefore,  until 
the  maximum  safe  charge  has  been  learned  from  experience. 
Good  results  will  often  be  had  in  hard  ground,  if  a  few 
cartridges  of  blasting  gelatin  are  used  in  the  point  of  the  boro 
hole,  with  gelatin  dynamite  on  top.  When  this  is  done,  it  is 
best  to  put  detonator  in  one  of  the  cartridges  of  blasting 
gelatin. 

Ammonia  dynamite  is  made  from  25  per  cent  to  75  per  cent 


442  COAL  MINING  COSTS 

strength.  All  grades  are  quicker  than  gelatin  dynamites,  and 
generally  speaking  the  quickness  increases  with  the  strength. 
That  is,  the  stronger  grades  are  quicker,  and  the  lower  grades 
slower,  in  action. 

The  various  grades  of  these  three  high  explosives,  offer  a 
wide  range  in  strength  and  quickness  to  select  from,  and  it  is 
always  possible  after  a  few  trials  to  find  an  explosive  exactly 
suited  to  the  conditions. 

Railroad  tunnels,  mine  tunnels  and  drifts,  highway  tunnels, 
and  irrigation  tunnels,  are  being  driven  daily  through  various 
kinds  of  "ground."  Often  it  is  a  matter  of  first  importance 
to  finish  them  quickly,  and  consequently  details  in  regard  to 
methods  and  equipment  are  matters  of  general  interest. 

In  Engineering  Contracting  of  October  20,  1909,  Mr.  J.  B. 
Lippincott,  assistant  chief  engineer  of  the  Los  Angeles  aqueduct, 
gave  an  interesting  account  of  the  driving  of  the  Red  Rock 
tunnel  of  the  Los  Angeles  aqueduct  system.  In  August,  1909, 
this  tunnel,  which  is  9-ft.  10  in.  X  10  ft-  Sy2  in.  in  section,  was 
advanced  1061.6  ft.  Mr.  Lippincott  states  that  the  explosives 
used  were  Du  Pont  40-per-cent  ammonia  dynamite  and  blasting 
powder. 

In  the  Engineering  News  of  November  18,  1909,  the  Red 
Rock  tunnel  is  again  referred  to,  and  details  are  also  given  by 
Mr.  C.  H.  Richards,  division  engineer,  in  regard  to  a  tunnel 
on  the  Little  Lake  Division  of  the  Los  Angeles  acqueduct.  The 
explosives  used  in  this  tunnel  were  Hercules  40-per-cent  and  60- 
per-cent  gelatin  dynamite,  the  average  weight  of  explosives  per 
cubic  yard  of  rock,  place  measurement,  having  been  only  3.3  lb., 
or  about  35  lb.  per  linear  yard  of  tunnel  almost  10  ft.  X  10  ft- 
in  section. 

A  short  time  before,  accounts  were  given  in  several  engineer- 
ing magazines,  of  a  record  driving  speed  made  in  the  Roosevelt 
drainage  tunnel,  at  Cripple  Creek,  Col.  The  explosives  used  in 
this  tunnel  were  40-,  50-,  and  60-per-cent  Repauno  gelatin  dyna- 
mite and  Du  Pont  blasting  gelatin. 

A  very  interesting  description  of  the  Rondout  pressure  tunnel 
of  the  Catskill  aqueduct,  written  by  John  P.  Hogan,  assistant 
engineer  of  the  New  York  City  Board  of  Water  Supply,  was 
published  in  the  January  1,  1910,  number  of  the  Engineering 
Record.  Very  rapid  progress  was  made  in  this  tunnel,  and  also 


MISCELLANEOUS  INSIDE  COSTS  443 

in  the  Moodna  pressure  tunnel  of  the  same  system,  described  in 
an  article  in  the  Engineering  Record  of  June  4,  1910.  The 
explosive  which  gave  best  results,  and  which  was  used  exclusively 
in  both  of  these  tunnels  was  60-per-cent  For  cite — a  gelatin  dyna- 
mite. 

Reference  to  a  paper  by  B.  H.  M.  Hewett  and  "W.  L.  Brown, 
on  the  land  section  of  the  Pennsylvania  Railroad  North  River 
tunnels,  published  in  Vol.  XXXVI  of  the  Proceedings  of  the 
American  Society  of  Civil  Engineers,  and  reprinted  in  part  in 
Engineering  Contracting  of  May  11,  1910,  shows  that  40-per- 
cent. Forcite  was  used  in  blasting  on  the  Manhattan  section, 
and  60-per-cent  Forcite  on  the  Weehawken  section. 

The  records  of  many  other  tunnels  recently  constructed, 
further  illustrate  how  many  kinds  and  strengths  of  explosives 
are  used  for  blasting  under  the  different  conditions  encoun- 
tered in  one  class  of  work. 

The  specific  cases  referred  to  above,  were  all  connected 
with  large  and  important  contracts,  where  equipment  and 
methods  were  at  the  best,  and  several  of  these  tunnels  were 
driven  at  record  speed.  The  fact  that  so  many  different 
explosives  were  used  in  the  several  tunnels,  goes  to  show  that 
care  was  taken  to  use  the  explosive  which  was  best  adapted 
to  the  conditions,  and  it  is  not  unlikely  that  the  speed  of  driv- 
ing these  tunnels,  was  largely  due  to  the  attention  given  to  the 
selection  of  the  explosives. 

VENTILATING  COSTS 

There  are  three  forms  of  efficiency  that  must  be  considered 
in  all  fans,  and  the  value  of  the  fan  depends  largely  upon  its 
ability  to  give  that  efficiency,  which  is  most  valuable  for  the 
particular  mine  on  which  the  fan  is  to  operate. 

The  mechanical  efficiency  is  to  be  considered  first.  Mechan- 
ical efficiency  is  the  relation  the  useful  work  performed  by  the 
fan,  bears  to  the  power  required  to  propel  the  fan.  The 
manometric  efficiency  is  of  second  importance.  Manometric 
efficiency  is  the  relation  which  the  depression  caused  by  the 
fan,  is  to  the  theoretic  depression  which  the  fan  would  make 
if  it  were  a  perfect  machine,  and  working  against  a  closed  air- 
way. The  volumetric  efficiency  is  the  third  to  be  considered. 


444  COAL  MINING  COSTS 

This  is  the  relation  which  the  volume  of  air  discharged  by  the 
fan,  bears  to  the  volume  or  cubic  contents  of  the  fan,  taken 
the  number  of  times  of  rotation  during  the  given  interval  of 
measurement. 

A  fan  may  be  high  in  one  of  these  efficiencies  and  low  in 
the  others.  For  instance,  a  fan  may  produce  a  large  volume 
of  air  at  a  very  low  pressure,  and  be  excellent  in  volumetric 
efficiency,  and  yet  be  low  in  manometric  efficiency  and  mechani- 
cal efficiency.  In  fact,  fans  high  in  mechanical  efficiency  are 
usually  high  in  manometric  efficiency,  and  almost  all  fans  high 
in  mechanical  efficiency,  when  working  under  favorable  con- 
ditions, are  low  in  volumetric  efficiency. 

A  fan  may  be  high  in  volumetric  efficiency  and  low  in 
mechanical  efficiency,  and  also  low  in  manometric  efficiency; 
these  conditions  all  depend  on  the  construction  of  the  fan,  and 
the  conditions  under  which  it  works.  The  mechanical  efficiency 
of  the  fan  may  be  determined  if  the  volume  of  air  passing 
through  the  mine  is  known,  and  the  pressure  (water-gage) 
reading  is  taken  at  the  same  time  the  air  reading  is  taken, 
together  with  the  indicated  power  of  the  motor  (either  steam 
engine  or  electric  motor)  when  resolved  into  foot-pounds. 

The  manometric  efficiency  may  be  determined  if  the  tangen- 
tial speed  of  the  fan  is  known,  and  the  actual  pressure  as 
read  by  the  water-gage.  The  volumetric  efficiency  may  be 
determined  if  the  rotations  of  the  fan  are  known,  and  the  vol- 
ume of  air  discharged  is  known  for  the  same  interval.  A  fan 
may  have  high  manometric  efficiency,  and  not  be  useful  for 
mine  ventilation;  for  instance,  a  cupola  fan  can  produce  the 
required  pressure  for  a  mine  and  not  give  any  volume  worth 
consideration.  A  fan  may  have  a  large  volumetric  efficiency 
under  favorable  conditions,  and  yet  under  unfavorable  con- 
ditions be  unable  to  produce  pressure,  and  hence  be  a  poor 
fan  for  mine  ventilation. 

Testing  a  fan  is  a  proceeding  requiring  a  considerable 
degree  of  skill  and  care.  It  is  customary  in  making  an  ane- 
mometer test  to  organize  the  force  so  that  each  main  airway 
shall  have  two  men  allotted,  one  to  take  the  anemometer  read- 
ings, and  one  to  hold  the  watch  and  light  and  call  the  time. 
The  time  of  readings  should  be  one  minute,  two  minutes,  three 
minutes,  or  any  number  of  minutes  that  may  be  agreed  upon. 


MISCELLANEOUS  INSIDE  COSTS  445 

It  is  difficult,  however,  to  hold  the  anemometer  in  a  swift  air 
current  longer  than  three  minutes,  and  for  this  reason  this 
is  usually  the  time  limit.  The  readings  should  be  taken  about 
100  ft.  away  from  the  fan  in  a  drift,  or  about  100  ft.  from  the 
bottom  of  the  shaft  in  a  shaft  mine,  and  should  be  taken  at  a 
place  where  the  section  of  the  airway  is  nearly  uniform  and  as 
smooth  as  possible,  and  not  close  to  any  turns,  where  the  air 
may  be  deflected  into  eddies. 

It  is  well  to  adopt  a  schedule  of  readings  at  intervals  of 
fifteen  or  twenty  minutes  so  that  the  velocity  of  air,  the  water- 
gage,  the  speed  of  motor,  and  temperature  and  barometer  may 
be  taken  at  exactly  the  same  time  as  the  motive  power  applied 
is  indicated.  With  large  fans  it  is  customary  to  use  steam 
engines,  and  the  indicator  is  used  to  ascertain  the  motive 
power  applied.  In  the  usual  tests  at  coal  mines,  the  barometer 
and  temperature  are  neglected,  as  not  having  sufficient  bear- 
ing to  warrant  the  calculations  necessary  to  apply  them  to 
the  test.  It  is  essential  that  a  certain  time  be  set  for  each 
and  every  reading,  that  they  may  be  all  taken  simultaneously, 
and  at  least  three  readings  should  be  taken  at  varying  speeds 
of  the  fan.  It  is  preferable  to  test  under  such  speeds  as  may 
be  employed  under  the  conditions  which  the  mine  may  require. 
To  get  proper  readings  of  the  water-gage,  it  should  be  so 
placed  that  it  will  record  the  pressure  of  the  air  immediately 
in  front  of  the  fan,  and  in  the  full  current  where  there  is  no 
possibility  of  an  eddy  in  the  current. 

To  obtain  the  proper  position  it  is  necessary  to  pipe  from 
the  water-gage  in  the  engine  room  to  a  place  in  front  of  the 
fan  where  the  air  current  is  in  one  body.  The  end  of  this  pipe 
should  be  from  10  to  25  ft.  from  the  periphery  of  the  fan  and 
should  be  bent  so  the  air  current  will  flow  directly  into  it. 
If  it  is  hung  down  a  shaft,  the  pipe  should  be  curved  returning 
upward,  and  have  a  small  hole  in  the  lower  side  of  the  curve 
to  drain  the  moisture  that  would  collect  and  form  a  water  seal, 
if  there  were  no  drainage.  This  water  seal  will  destroy  the 
true  reading  if  the  pipe  is  not  drained. 

The  indicator  should  be  operated  by  a  competent  engineer 
who  understands  the  theory  of  steam  consumption  of  a  steam 
engine  as  well  as  the  mere  knowledge  of  the  operating  of  a 
steam-engine  indicator,  as  there  may  be  defects  in  the  valve 


446 


COAL  MINING  COSTS 


gear  that  the  indicator  will  show  to  the  practiced  eye,  which 
might  be  overlooked  by  the  novice.  It  should  be  borne  in 
mind  that  the  engine  or  motor  is  having  its  efficiency  tested 
along  with  the  fan,  and  any  loss  of  efficiency  in  the  engine  or 
motor  will  reflect  on  the  fan.  In  taking  the  readings  it  is  well 
to  begin  with  the  even  hour,  and  allow  five  minutes  to 
each  party  detailed  to  take  the  various  readings  to  measure  the 
air  velocity,  the  water-gage  and  to  take  the  indicator  dia- 
grams. The  fan  should  be  maintained  at  a  constant  speed 
during  this  interval.  At  the  end  of  the  first  five  minutes,  the 
fan  should  be  accelerated  to  a  desired  speed  during  the  succeed- 
ing ten  minutes,  and  beginning  with  the  even  quarter  hour, 
be  maintained  at  a  constant  speed  during  the  succeeding  five 
minutes,  so  the  various  readings  may  be  taken  at  the  desired 
speed  of  the  fan.  At  the  end  of  this  five  minutes,  the  fan  may 
be  accelerated  again,  and  other  readings  made  at  intervals  of 
one  quarter  hour,  and  as  long  as  may  be  desired. 

The  following  table  gives  a  test  taken  by  the  West- 
moreland Coal  Company,  of  Irwin,  Penn.,  which  is  a  fair  sample 
of  the  manner  of  taking  the  various  readings  on  a  20-ft.  cen- 
trifugal fan  driven  by  a  steam  engine: 


FAN  TEST  AT  WESTMORELAND  SHAFT,  IRWIN,  PA. 


SPLIT  No.  1 

SPLIT  No.  2 

SPLIT  No.  3 

Time. 
P.M. 

per 

Volume 

W.  G. 

H.P. 
in  Air 

H.P. 
Ind. 

Mech. 
Eff. 

Area 

Vel. 

Area 

Vel. 

Area 

Vel. 

4  :  00 

100 

50 

1428 

58 

1593 

44 

500 

195,134 

2.40 

73.8 

114 

65 

4  :  15 

132 

50 

2245 

58 

2200 

44 

740 

286,791 

4.20 

190 

265 

71.7 

4  :  30 

150 

50 

2853 

58 

2383 

44 

822 

335,914 

5.10 

272 

344 

79.0 

4  :  45 

150 

50 

2800 

58 

2463 

44 

833 

337,171 

5.20 

276 

360 

77.7 

The  table  shows  the  general  method  of  a  test  for  practical 
purposes,  where  boiler  test  is  omitted,  and  also  the  variation 
of  the  barometer.  For  all  practical  purposes,  this  test  is  suffi- 
cient, and  shows  the  mine  engineer  where  he  can  practice 
economy  on  his  ventilation,  through  the  mechanical  efficiency 
of  his  fan. 

Power  required. — The  power  required  to  drive  air  through 
a  given  mine  increases  as  the  cube  of  the  volume,  provided  no 


MISCELLANEOUS  INSIDE  COSTS  447 

change  is  made  in  the  air  courses:  Let  V  represent  volume  of 
air,  in  cubic  feet  of  air  per  minute ;  let  P  represent  pressure  of 
air  as  represented  by  inches  in  water-gage;  Let  W  represent 
pressure  in  foot-pound  performed  on  the  air ;  then  W  =  V  X  I" 
X  5.2,  as  5.21  Ib.  is  the  pressure  of  air  per  sq.ft.  as  represented 
by  1  in.  of  water-gage.  When  reduced  to  horsepower,  the 
formula  then  becomes 

FXPX5.2 


h.p. 


33,000 


To  illustrate  this  formula,  suppose  a  mine  passing  100,000 
cu.ft.  of  air  per  min.  against  a  2.3-in.  water-gage.  The  work 
performed  in  passing  the  quantity  of  air  at  the  given  pressure, 
will  be  in  terms  of 

100,000X2.3X5.2 


hp.= 


33,000 


which  is  equal  to  36.24  hp.  Now  the  pressure  of  air  increases 
as  the  square  of  the  volume,  so  that  if  we  desire  to  increase 
the  volume  of  air  from  100,000  cu.ft.  of  air  per  min.  in  the 
above  instance  to  200,000  cu.ft.  of  air  per  min.,  we  will  re- 
quire eight  times  as  much  power  as  to  produce  the  100,000 
cu.ft.  of  air  per  min.,  as  in  the  first  case,  for  having  doubled 
the  velocity,  the  water-gage  will  have  increased  from  2.3  in. 
to  four  times  2.3,  or  9.2  in.  Now  substituting  in  our  second 
formula,  we  have 

,     _  200,000X9.2X5.2 
P'  33,000 

which  equals  289.92  hp.,  which  is  eight  times  36.24  hp.,  the 
amount  of  power  required  in  the  first  example.  It  will  be 
seen  that  to  get  the  increased  100,000  cu.ft.  of  air,  it  will 
require  252.68  hp.,  whereas  in  the  first  example  only  36.24  hp. 
was  required,  so  that  the  second  100,000  cu.ft.  of  air  required 
seven  times  as  much  power  to  pass  it  through  the  mine  as  the 
first  100,000  cu.ft. 

When  we  consider  that  the  installation  of  a  ventilator, 
capable  of  producing  289  hp.  in  the  air,  would  cost  eight  times 
as  much  as  a  ventilator  producing  36  hp.  in  the  air,  we  can 
understand  why  the  management  hesitates  to  purchase  such 


448  COAL  MINING  COSTS 

expensive  apparatus.  Furthermore,  if  a  horse-power  in  the 
most  favored  coal  regions  at  the  mine  costs  about  $50  per 
year,  then  to  increase  the  ventilation  as  above  cited  from  100,- 
000  cu.ft.  of  air  per  min.  to  200,000  cu.ft.  of  air  per  min.,  would 
cost  for  power  alone,  about  $12,634  per  year.  This  annual  cost 
at  many  mines  would  be  a  serious  consideration.  What,  then, 
would  be  required  to  get  the  necessary  amount  of  air  without 
making  such  an  outlay  of  money  in  running  expense?  In 
most  mines  the  quantity  of  air  could  be  doubled  without  in- 
creasing the  power  eight  times,  by  cleaning  up,  enlarging 
and  increasing  the  number  of  air  courses,  thereby  reducing 
the  velocity  of  air  and  consequently  the  pressure.  The  pres- 
sure is  the  great  absorber  of  power,  and  it  is  well  to  bear  in 
mind  that  for  a  given  amount  of  air,  a  certain  pressure  will  be 
necessary  to  propel  the  air  through  the  mine,  and  regardless 
of  what  form  of  ventilator  is  used,  this  pressure  will  be  the 
same. 

It  is  a  common  but  mistaken  belief  among  mining  men,  that 
one  favored  form  of  ventilator  will  propel  the  current  through 
a  mine  at  a  less  pressure  than  another  ventilator.  This  fallacy 
should  not  be  entertained  by  any  mine  manager. 

Specification  of  fans. — The  engineer  is  sometimes  called 
upon  to  prepare  specifications  for  a  fan  to  be  used  at  a  par- 
ticular mine  and  to  obtain  bids  for  the  erection  of  same.  Cer- 
tain requirements  will  be  laid  down,,  say  that  the  fan  will 
furnish  250,000  cu.ft.  per  minute  against  a  6-in.  water  gage. 
A  manufacturer  in  presenting  his  bid  will  sometimes  state  that 
he  can  do  better  than  what  the  specifications  ask,  often  claim- 
ing that  the  dimensions  given  are  too  great  and  the  fan  un- 
necessarily large.  Where  the  manufacturer  is  left  to  make  his 
own  estimate  of  the  size  of  fan  required,  he  will  often  make 
no  inquiries  in  reference  to  the  size  of  the  mine  airways  or  the 
present  circulation  in  the  mine ;  but  will  take  chances  in  refer- 
ence to  these  important  data.  In  nine  cases  out  of  ten  the 
new  fan  is  installed  at  a  new  mine,  where,  owing  to  the  short 
airways  of  ample  size,  it  continues  to  do  good  work  for  a  few 
years.  In  the  later  development  of  the  mine,  however,  there  is 
experienced  a  scarcity  of  air,  and,  referring  to  the  guaranteed 
capacity  of  the  fan,  the  superintendent  orders  the  mine  fore- 
man to  speed  it  up,  claiming  that  it  is  not  run  at  its  full 


MISCELLANEOUS  INSIDE  COSTS  449 

capacity.  The  results  are  still  far  from  satisfactory,  but  the 
fault  is  not  with  the  fan,  which  is  now  forced  to  operate  under 
conditions  for  which  it  was  not  designed. 

A  short  familiar  calculation  of  the  pressure  required  to  pass 
a  given  quantity  of  air  through  an  airway  of  given  dimensions, 
using  the  Fairley  coefficient  (k  =  0.00000001) ,  shows  that  a  water 
gage  of  1.9  in.  is  required  to  pass  100,000  cu.ft.  of  air  per 
minute  through  an  airway  7  X  10  ft.  in  section,  for  each  1000 
ft.  in  length.  Suppose  the  main  airway  in  the  mine  is  3000  ft. 
long  to  the  first  point  of  split.  The  distance  the  main  air  cur- 
rent must  traverse,  including  the  return,  is  then  6000  ft.,  and 
the  water  gage  required,  for  the  passage  of  the  main  airways 
only,  is  6  X  1-9  —  H-4  in.  It  is  clear  at  once  to  any  mining 
engineer  or  foreman  that  it  would  be  impracticable  to  attempt 
to  pass  100,000  cu.ft.  of  air  through  a  single  airway  of  this 
size  for  a  distance  of  6000  ft.,  including  the  return.  By  pro- 
viding two  air-ways  of  this  size  for  the  intake  and  two  air- 
ways for  the  return,  both  the  water  gage  and  the  power  on 
the  air  would  be  reduced  to  one-fourth  of  the  previous  amount. 

By  this  means,  a  main  air  current  of  100,000  cu.ft.  per  min. 
can  be  conducted  a  distance  of  3000  ft.  to  the  first  point  of 
split  and  returned  to  the  upcast,  under  a  water  gage  of  2.85 
in.,  requiring  45  hp.  This  is  the  horsepower  on  the  air  con- 
sumed in  the  main  airway.  These  are  the  more  important 
calculations  in  estimating  the  requirements  of  the  mine  in 
respect  to  ventilation.  To  these  amounts,  however,  must  be 
added  the  water  gage  and  power  consumed  in  the  splits. 

Finally,  to  determine  the  power  of  the  engine  driving  the 
ventilating  fan,  it  is  customary  to  assume  an  efficiency  of  60 
per  cent.  For  example,  if  the  entire  horsepower  of  the  air, 
for  the  main  airways  and  splits,  is,  say  60  hp.,  the  required 
horsepower  of  the  engine  would  be  60  -r-  0.60  =  100  hp.  A 
similar  calculation  shows  that  a  water  gage  of  5.08  in.  and 
160  hp.  on  the  air  are  required  to  pass  200,000  cu.ft.  per  min. 
in  three  airways,  a  distance  of  6000  ft. ;  but,  using  four  air- 
ways, the  same  air  volume,  circulated  the  same  distance,  will 
only  require  a  water  gage  of  2.85  in.  and  90  hp.  on  the  air. 

Assuming  a  consumption  of  5  Ib.  of  coal  per  hp.-hr.,  the 
saving  of  160  —  90  =  70  hp.,  by  employing  four  instead  of 
three  main  airways,  would  correspond  to  a  saving  in  fuel  of 


450  COAL  MINING  COSTS 

5  X  70  =  350  Ib.  of  coal  per  hour,  or  1%  tons  in  10  hr.,  whicl) 
figures  are  a  fair  approximation  to  fact,  in  actual  practice. 
Assuming,  then  a  saving  of  3  tons  of  coal  per  day  of  24  hr.  (less 
fuel  being  consumed  during  the  night,  when  the  mine  is  not 
working),  at  a  cost  of  $1  per  ton  for  coal  at  the  mines,  the  total 
saving  per  month  would  be  practically  3  X  30  =  90  tons  or 
$90  per  month,  at  the  least  computation. 

Experience  suggests  that  economical  ventilation  requires 
that  the  water  gage  be  kept  down  to  at  least  1  in.  per  100,000 
cu.ft.  of  air  in  circulation.  As  has  often  been  suggested  before, 
this  can  be  done  by  properly  splitting  the  air  current  in  the 
mines.  Experience  indicates  further,  that  for  the  circulation 
of  from  80,000  to  100,000  cu.ft.  of  air  per  minute,  two  main 
intakes  and  two  return  airways  should  be  provided;  or  three 
main  airways  for  a  circulation  up  to  150,000  cu.ft.  per  min. ; 
and  four  main  airways  for  a  circulation  up  to  200,000  cu.ft. 
per  min.,  the  sectional  area  being  70  sq.ft.  If  this  plan  is 
followed,  a  great  saving  in  coal  will  be  found  to  result.  It  is 
important,  in  overcasting  the  air,  to  see  that  all  air  bridges 
on  main  airways  are  substantially  built  so  as  to  prevent  the 
leakage  of  air.  Such  overcasts  and  all  stoppings  on  main  air- 
ways should  be  built  of  concrete,  for  the  same  reason. 

It  has  been  the  policy  of  the  Consolidation  Coal  Co.,  when 
opening  a  new  mine,  to  make  the  projections  on  the  maps  of 
that  mine,  with  enough  intakes,  returns  and  overcasts,  to 
ventilate  the  mine  under  a  low  water  gage.  Where  old  mines 
were  contending  with  a  high  water  gage,  air  shafts  were  sunk, 
in  order  to  shorten  the  airways  and  lower  the  water  gage. 
At  Mine  No.  26,  an  air  shaft  was  sunk  11,700  ft.  from  the 
mouth  of  the  mine,  and  a  new  6  X  20-ft.  fan  and  a  boiler  planl 
installed.  By  installing  the  fan  at  this  last  location,  it  shortened 
the  airway  one-half,  besides  permitting  the  two  return  airways 
to  be  used  as  outlets  or  intakes,  as  required,  thus  providing 
four  inlets  in  place  of  two,  as  before.  Previous  to  the  change, 
this  mine  had  but  two  inlets  and  two  returns.  By  moving  the 
fan,  the  water  gage  was  reduced  from  3  in.  for  100,000  cu.ft.. 
to  1.36  in.  for  the  same  air  volume,  a  splendid  demonstration 
of  the  practical  value  of  shortening  airways.  It  will  be  noticed 
that  1.36  in.  water  gage  is  too  high  for  a  circulation  of  100,000 
cu.ft.  of  air,  as,  in  case  200,000  cu.ft.  were  required  the  neces- 


MISCELLANEOUS  INSIDE  COSTS  451 

sary  gage  would  be  5.44  in.,  which  is  far  too  high  for  every- 
day practice.  It  may  be  possible  that  this  is  the  best  that  can 
be  done  with  an  old  mine  sometimes. 

In  estimating  the  saving  of  coal  effected  by  moving  this 
fan,  the  calculation  is  based  on  the  difference  in  actual  total 
horsepower  required  of  each  engine,  which  is  123  hp.  for  150,- 
000  cu.ft.  of  air.  Five  pounds  of  coal  per  hp.  per  hr.  then 
means  615  Ib.  of  coal  per  hr.,  or  14,760  lb.,  or,  say  7.4  net  tons 
per  day  of  24  hr.  Coal  at  $1  per  ton  then  means  $7.40  per  day. 
or  $222  per  month  and  $2664  per  year.  The  water  gage  on 
the  new  fan  will  be  3.06  in.  for  150,000  cu.ft.  of  air ;  while  on 
the  old  fan,  it  would  require  a  6.75  in.  water  gage  for  the 
same  air  volume. 

The  Consolidation  Coal  Co.  has  adopted  a  different  method 
for  the  ventilation  of  its  new  coal  field  in  the  Elkhorn  Division, 
Kentucky.  The  mines  are  all  projected  on  the  map  of  the  coal 
field,  showing  the  mines  about  as  they  will  be,  and  giving  the 
amount  of  coal  they  expect  each  mine  to  produce  and  the 
amount  of  air  they  expect  to  put  into  each  mine.  Then  the 
proper  number  of  airways  are  projected  on  the  map,  together 
with  all  overcasts,  so  as  to  handle  this  amount  of  air  with  a 
given  water  gage.  The  large  mines  from  which  it  is  intended 
to  produce  not  less  than  2000  tons  of  coal  per  day,  are  expected 
to  require  about  200,000  cu.ft.  of  air  per  minute  when  the 
mine  becomes  fully  developed. 

In  these  large  mines,  there  will  be  four  return  airways  and 
four  intakes,  each  of  which  will  average  6^2  ft.  high  by  10  ft. 
wide,  making  an  area  of  65  sq.ft.  each.  From  the  fan  to  the 
first  split  is  generally  about  1200  ft.,  and  here  two  double- 
face  headings  are  turned  off — two  to  the  right  at  about  1200 
ft.  and  two  to  the  left  at  about  1320  ft. — so  the  switches  will 
not  interfere,  as  the  face  headings  are  turned  with  a  112-ft. 
radius  curve.  Each  of  these  two  face  headings  will  have  about 
four  room  headings,  making  eight  room  headings  for  each 
25,000  cu.ft.  of  air.  The  four  room  headings  will  be  ventilated 
with  one  split  of  air,  say  about  12,500  cu.ft.  to  each  four  room 
headings,  which  will  be  about  1000  ft.  long.  Face  headings 
will  be  turned  every  2100  ft.,  center  to  center,  and  room  head 
ings  are  then  to  be  turned  both  ways  from  the  face  heading. 

The  water  gage  that  will  be  required  in  passing  200,OOG 


452  COAL  MINING  COSTS 

GU.it.  of  air  to  the  first,  second,  third  and  fourth  splits,  there 
being  four  return  airways  and  four  intake  airways,  is  esti- 
mated as  follows:  The  distance  to  the  first  split  of  air  is,  say 
1260  ft.,  an  average  between  the  two  face  headings  where 
50,000  cu.ft.  of  air  goes  to  the  four  face  headings.  Using 
Fairley's  coefficient  of  friction  (k  =  0.00000001),  the  water  gage 
for  200,000  cu.ft.  of  air  up  to  this  first  split  will  be 

0.00000001  X  1260X2(6.5+10)  X200,0002     n  _0  . 
W'g=  5.2(6.5X10)3  =(X72m- 

To  the  second  split  is  a  distance  of  2100  ft.,  the  air  volume 
being  150,000  cu.ft.,  the  water  gage  for  this  section  is  found 
in  the  same  manner,  and  is  0.68  in.  Likewise,  to  pass  100,000 
cu.  ft.  of  air  to  the  third  split,  a  distance  of  2100  ft.  again, 
will  require  a  water  gage  of  0.30  in.,  and  finally,  to  pass  the 
remaining  50,000  cu.  ft.  the  same  distance  to  the  last  face 
headings  will  require  a  gage  of  0.075  in.  The  sum  of  these 
four  gages  must  then  be  doubled  to  provide  for  the  return, 
in  each  case,  which  gives  for  the  total  water  gage  absorbed 
in  the  main  headings  1.775  X  2  =  3.55  in. 

It  is  well  to  estimate  on  a  loss  of  at  least  10  per  cent  of  the 
air,  which  means  a  loss  in  water  gage  of  19  per  cent,  which 
taken  from  3.55  in.  leaves  only  2.87  in.  water  gage  for  200,000 
cu.ft.  of  air  on  the  whole  mine,  to  this  point.  The  splits  in  the 
face  headings  show  a  water  gage  of  only  0.075  in.  for  2100  ft. 
with  12,500  cu.ft.  of  air. 

All  regulators  on  face  headings  should  be  placed  near  the 
main  airway  and  must  not  be  set  beyond  the  first  room  head- 
ing, under  any  consideration.  In  reference  to  the  percentage 
of  water  gage  spoken  of  as  being  lost  (19  per  cent),  it  has 
been  found  in  practice  this  loss  reaches  as  high  as  25  per  cent, 
so  that  it  is  safe  to  call  it  19  per  cent,  as  10  per  cent  loss  in 
air  means  19  per  cent  loss  in  gage.  This  is  not  all  due  to  loss 
of  air  by  leakage,  but  from  air  passing  crosscuts  and  other 
wide  places,  which  reduce  the  velocity. 

Most  of  the  old  style  paddle-wheel  and  screw-propeller  fans 
are  working  with  less  than  20  per  cent  mechanical  efficiency, 
while  the  latest  high-grade  speed  fans  are  working  above  60 
per  cent  mechanical  efficiency,  and  in  the  great  majority  of 
cases,  are  working  with  a  mechanical  efficiency  of  between  70 


MISCELLANEOUS  INSIDE  COSTS  453 

and  80  per  cent.  Assuming  that  the  old-style  fan  is  working  at 
20  per  cent  mechanical  efficiency,  and  the  new  high-speed  fan 
at  60  per  cent  efficiency,  also  that  100,000  cu.ft.  of  air  per  min. 
at  a  2-in.  water-gage  is  required,  the  horse-power  in  air 

100,000X2X5.2 

33,000         =31-51^' 

Now  31.51  hp.  is  20  per  cent  of  157  hp. ;  31.51  hp.  is  60  per 
cent  of  52.51  hp.  It  will  be  seen  that  157.55  hp.  will  be  required 
to  drive  the  old-style  fan  to  produce  the  required  ventilation, 
while  only  52.51  hp.  will  be  required  to  drive  the  new  high- 
speed fan.  Thus  the  new  high-speed  fan  will  save  the  dif- 
ference between  157.55  hp.  and  52.51  hp.,  or  105.04  hp.  Now, 
as  one  horse-power  is  produced  at  a  cost  of  $50  per  year,  it 
follows  that  the  saving  effected  with  the  new  high-speed  fan 
is  105.04  X  50  =  $5252  per  year.  This  amount  of  money  will 
install  a  fan  to  produce  the  required  100,000  cu.ft.  of  air  per 
min.  at  a  2-in.  water-gage,  under  the  worst  conditions  likely 
to  be  found,  and  would  therefore  return  to  the  owner  the 
entire  value  of  the  fan  each  year. 

Change  in  volume  of  air  required. — Ventilating  conditions 
at  all  mines  change  from  week  to  week.  As  the  workings 
extend  underground,  the  amount  of  air  required  is  increased, 
the  resistance  runs  up,  the  number  of  stoppings  multiply,  the 
current  of  air  may  be  lengthened  or  shortened.  All  of  these 
things  take  place,  resulting  in  a  change  in  the  amount  of  work 
required  of  the  fan.  If,  by  making  repairs,  extensions  and 
changes,  costing,  for  example,  $100,  the  fan  is  relieved  of  10 
per  cent  of  its  work,  we  have  something  to  show  on  both  sides 
of  the  account,  and  can  decide  whether  or  not  the  money  has 
been  well  spent,  if  we  know  the  amount  of  power  saved  and 
its  cost  per  unit. 

There  are  few  cases  where  the  cost  of  power  per  horse- 
power-hour is  known  in  installations  where  a  steam  engine  is 
used  for  driving  the  fan.  The  practice  usually  is  to  operate 
at  a  certain  number  of  revolutions  per  minute,  this  speed  being 
maintained  by  opening  and  closing  the  throttle  valve.  This 
critical  speed  is  governed  by  the  needs  of  the  mine.  If  there 
is  "bad  air"  on  "Third  Right,"  the  foreman  may  " speed 
'er  up"  a  few  revolutions.  Or,  again,  the  drivers  are  "not 


454  COAL  MINING  COSTS 

able  to  keep  a  light,"  in  another  part  of  the  mine.  In  this 
case  the  fan  may  be  ' '  slowed  down  a  little. ' ' 

Under  such  conditions,  and  they  exist  more  generally  than 
one  would  imagine,  the  cost  of  power  used  cannot  always  be 
figured  accurately.  In  many  instances  it  is  not  known  at  all. 
The  mine,  however,  may  be  amply  ventilated  at  some  one  of 
the  different  speeds  at  which  the  fan  is  run.  This  particular 
speed,  whatever  it  is,  is  the  proper  one  at  which  the  fan  should 
be  driven  during  the  time  when  the  underground  operations 
are  in  full  activity. 

Where  a  mine  is  equipped  with  a  fan  of  ample  capacity, 
and  has  airways  as  well  as  underground  structures  in  fair  con- 
dition, a  certain  number  of  revolutions  of  the  fan  will  furnish 
a  sufficient  amount  of  air  to  comply  with  the  mining  law  and 
to  fully  ventilate  the  mine  when  it  is  producing  its  maximum 
tonnage.  Such  ventilation  will  require  the  maximum  horse- 
power applied  to  the  shaft  of  the  fan. 

During  the  hours  when  the  mine  is  not  producing  coal,  say 
at  night,  on  Sundays  and  on  holidays,  only  a  fraction  of  this 
ventilation  is  needed.  In  mines  where  no  inflammable  gas  is 
found,  tests  have  shown  that  at  night  and  on  idle  days,  about 
half  the  full  ventilating  current  is  sufficient  for  all  needs  and 
all  the  demands  for  safety. 

Under  conditions  such  as  we  have  outlined,  a  steam-driven 
fan  can  be  slowed  down  to,  say,  half  the  speed  used  in  tKe 
daytime,  and  the  volume  of  air  will  be  reduced  in  almost  the 
same  proportion.  The  steam,  however,  will  vary  in  pressiire, 
and  this,  too,  will  further  vary  the  speed  of  the  fan  as  well 
as  the  volume  of  air  passing. 

In  nongaseous  mines  and  others  where  only  half  the  normal 
ventilation  is  required  when  the  mine  is  idle,  it  is  an  inexcus- 
able waste  to  use  more.  The  responsible  officials  at  the  mine 
should  decide  what  amount  of  air  is  necessary  when  the  mine 
is  not  working,  making  due  allowance  for  any  men  who  may 
be  required  in  the  mine  at  such  times  and  this  should  be  rigidly 
adhered  to. 

If  a  speed-regulating  device  were  supplied  as  in  the  case 
of  a  steam  engine,  for  example,  the  control  of  the  power  con- 
sumption, and  therefore  to  a  certain  degree  the  cost  of  ventila- 
tion, would  be  in  the  hands  of  a  comparatively  irresponsible 


MISCELLANEOUS  INSIDE  COSTS  455 

attendant  who  cannot  be  depended  on  to  do  one  thing  the 
same  way  every  time.  A  motor  capable  of  running  at  full 
speed,  whatever  that  may  be,  and  at  half  speed  fills  the  require- 
ments of  fan  drive  fully. 

Once  it  is  installed  it  assumes  all  responsibility  as  to  the 
speed  of  the  ventilating  apparatus.  It  operates  at  either  one 
or  the  other  of  its  two  possible  speeds,  and  the  fan,  as  a  result, 
is  either  handling  its  full  prescribed  capacity  or  approximately 
half  of  it. 

Occasions  arise  which  demand  that  the  maximum  rotation 
of  the  fan  be  changed,  either  increased  or  decreased.  Such 
changes  can  be  well  and  cheaply  made  by  substituting  a  dif- 
ferent diameter  of  pulley  on  the  motor  shaft,  when  a  belted 
installation  is  being  considered. 

The  power  saved  by  making  the  high-speed  rotation  exactly 
what  is  needed  to  fully  ventilate  the  mine,  and  no  more,  when 
underground  conditions  are  reasonably  good,  soon  amounts  to 
enough  to  pay  for  an  extra  or  different  driving  pulley,  or  even 
several  of  them. 

Money  can  frequently  be  saved  and  better  mine  ventilation 
assured  by  the  use  of  canvas  piping.  Wherever  blind  entries, 
such  as  water  courses,  air  courses,  tunnels  and  shafts  are 
contemplated  or  wherever  gases  will  not  clear,  this  system 
can  be  used  to  an  advantage.  Every  mine  should  carry  at 
least  one  of  these  outfits  on  hand  all  the  time. 

Briefly,  the  system  consists  of  a  flexible,  pliable,  treated 
canvas  tubing  and  a  fan  directly  mounted  on  the  armature 
shaft  of  a  motor.  Protection  for  the  tubing  is  afforded  by 
suspending  it  from  a  wire  attached  to  pegs  driven  in  the  roof 
at  15-ft.  intervals,  or  to  a  wire  fastened  to  upright  posts.  Addi- 
tional sections  are  coupled  within  15  seconds  with  a  special 
coupling  furnished  as  an  integral  part  of  each  section. 

The  outfit  provides  a  compact,  portable,  practical,  "fool- 
proof" blowing  unit,  to  be  used  as  an  auxiliary  to  the  large 
mine  fan.  In  nearly  all  mines,  places  occur  where  the  air  has 
been  short-circuited  to  such  an  extent  it  does  not  provide  suffi- 
cient ventilation  to  carry  off  gases  and  smoke  in  blind  head- 
ings. By  installing  a  booster  outfit  at  this  point,  the  air  can 
be  carried  for  distances  up  to  1000  ft.  in  blind  entries  in  suffi- 
cient volume  to  work  up  to  eight  men.  This  added  circulation 


456  COAL  MINING  COSTS 

of  air  strengthens  the  main  air  system  to  permit  work  of  any 
class  without  the  slightest  handicap  or  loss  of  time. 

It  is  often  possible  to  drive  headings  25  per  cent  faster  by 
eliminating  cross-cuts,  and  save  $50  to  $100  through  the  elimi- 
nation of  each  stopping.  These  units  are  capable  of  giving 
volumes  of  air  varying  from  1900  to  1200  cu.ft.  per  minute  for 
distances  from  100  to  1000  ft.  and  are  especially  designed  for 
driving  single  entries  in  coal  mines. 

Spiral  riveted  galvanized  pipe  is  also  often  used  in  ventilat- 
ing tunnels  and  gangways  in  which  there  are  no  return  air- 
ways. A  system  in  general  use  is  to  place  an  electric  booster 
fan  to  drive  air  through  an  18-in.  pipe  to  the  face  of  the  tunnel 
or  gangway,  and  allow  the  air  to  return  out  the  tunnel  or 
gangway,  until  a  hole  is  driven  to  the  surface  or  the  upper 
level,  when  the  fan  is  moved  nearer  the  face  and  pipe  used 
again  to  advance  the  tunnel  or  gangway.  Gangways  and 
tunnels  four  miles  long  have  been  driven  by  this  method ;  they 
are  usually  ventilated  a  distance  of  2000  ft.,  when  fan  must 
be  advanced. 

Mine  lighting. — To  measure  accurately  the  illuminating 
power  of  a  lamp,  we  must  consider  not  only  the  intensity  of 
light  (or  candlepower),  but  also  the  solid  angle  over  which 
the  intensity  is  maintained.  A  lamp  which  gives  an  intensity 
of  light  of  one  candlepower  all  around  gives  twice  as  much 
light  as  one  which  gives  a  light  of  equal  intensity  half  way 
round  it.  The  term  "flux"  is  used  by  illuminating  engineers 
to  designate  the  product  of  intensity  and  the  angle  over  which 
it  is  exhibited,  since  this  product  most  represents  the  light 
which  flows  from  the  lamp.  The  unit  of  flux  is  called  a  lumen 
and  is  about  8/100  of  the  total  flux  of  light  produced  by  a 
source  of  one  spherical  candlepower. 

The  term  candlepower  used  without  qualification  is  not 
only  confusing  but  really  meaningless.  If  all  sources  of  light 
distributed  light  equally  in  all  directions,  then  a  single  measure- 
ment of  their  candlepower  would  suffice  to  compare  them. 
Practically,  however,  sources  of  light  differ  a  great  deal  in  the 
way  they  distribute  light,  and  this  is  especially  true  if  re- 
flectors are  used. 

Therefore,  if  a  lamp  is  stated  to  give  two  candlepower,  the 
statement  should  also  explain  whether  "  head-on "  candlepower 


MISCELLANEOUS  INSIDE  COSTS  457 

is  meant,  or  average  candlepower  over  the  stream  of  light,  or 
average  candlepower  in  a  given  plane — such  as,  for  instance, 
the  horizontal.  A  lamp  that  uses  a  reflector  may  have  a  *  *  head- 
on"  candlepower  3  to  10  times  the  average  candlepower  over 
its  entire  stream  of  light.  Generally  it  is  best  to  state  the 
average  candlepower  of  a  lamp  instead  of  the  candlepower 
at  a  single  point  or  group  of  points. 

A  statement  of  the  candlepower  of  a  lamp  does  not  suffi- 
ciently define  its  light-giving  capacity.  A  100-cp.  lamp  is 
seemingly  33  times  as  desirable  as  a  3-cp.  lamp  and  yet  a 
100-cp.  lamp  shining  through  a  hole  %  in.  in  diameter  gives 
less  actual  light  and  much  less  useful  light  than  a  3-cp.  lamp 
shining  through  a  hole  3  in.  in  diameter.  Therefore,  in  order 
to  define  properly  the  light-giving  capacity  of  a  lamp,  a  state- 
ment must  be  made  regarding  both  the  candlepower  and  the 
total  flux  of  light  (or  lumens)  produced  by  the  lamp. 

The  selection  of  proper  lower  limits  for  intensity  of  light 
and  its  flux  is,  aside  from  safety,  the  most  important  con- 
sideration in  selecting  portable  electric  lamps.  Without  these 
standards  of  reference  accurate  and  intelligent  comparison  of 
lamps  is  not  possible.  In  an  attempt  to  establish  such  lower 
limits  the  Bureau  of  Mines  searched  for  some  time  for  stand- 
ards which  should  be  fair,  not  too  low  in  value,  not  arbitrarily 
selected,  and  which  should  bear  an  easily  recognized  relation  to 
something  already  in  use. 

It  was  finally  decided  to  prepare  a  standard  Wolf  safety 
lamp  to  give  its  best  performance  and,  after  adjusting  the 
flame  height  to  1  in.,  measure  the  average  intensity  of  the 
stream  of  light  and  also  the  total  flux  of  light  in  the  stream. 
This  was  accordingly  done  at  two  different  times,  using  dif- 
ferent lamps,  prepared  by  different  men,  and  tested  with  dif- 
ferent instruments  of  different  types.  The  first  measurements 
were  made  by  Dr.  L.  0.  Grondahl,  of  the  Carnegie  Institute  of 
Technology,  and  the  second  measurements  at  the  Bureau  of 
Mines.  The  results  of  the  two  tests  checked  within  a  very  few 
per  cent. 

The  lamp  used  was  a  Wolf  miner's  safety  lamp,  1907  model, 
round  burner,  burning  70-72  deg.  naphtha,  and  prepared  and 
trimmed  in  accordance  with  the  standard  practice  of  the  Bureau 
of  Mines.  The  average  intensity  of  light  stream,  as  determined 


458  COAL  MINING  COSTS 

by  these  tests,  was  a  trifle  under  0.4  cp.  and  the  total  flux  of 
light  was  found  to  be  not  quite  3  lumens. 

The  bureau  therefore  concluded  that  a  satisfactory  lower 
limit  of  flux  of  light  for  hand  lamps  would  be  3.0  lumens  and 
a  satisfactory  lower  limit  of  average  intensity  would  be  0.4 
candlepower. 

The  bureau  suggests  that  lamps  designed  to  be  worn  upon 
the  cap  should  give  the  same  intensity  of  light  as  that  required 
for  hand  lamps,  but  that  the  minimum  flux  of  light  required 
from  cap  lamps  should  be  not  more  than  half  the  minimum 
demanded  from  hand  lamps,  because  when  a  lamp  is  worn 
upon  the  head  any  light  that  is  thrown  to  the  rear  is  wasted. 
If  the  equivalent  of  a  safety  lamp  were  mounted  upon  a  man's 
head,  one-half  of  its  light  would  fall  behind  the  man  and  thus 
could  not  be  used.  Therefore  the  bureau  concluded  that  1.5 
lumens  would  be  a  satisfactory  lower  limit  for  the  flux  of  light 
produced  by  a  cap  lamp. 

Twelve  hours  was  selected  by  the  bureau  as  a  reasonable 
time  of  burning.  This  length  of  time  was  chosen  after  con- 
sultation with  several  people  outside  of  the  bureau,  who  were 
competent  to  express  an  opinion  in  regard  to  the  subject. 

The  folowing  table  prepared  by  E.  M.  Chance,  gives  com- 
parative candlepower  of  various  types  of  lamps.  These  data 
were  accumulated  during  about  eight  years  and  are  general 
averages.  The  photometric  determinations  were  made  upon 
a  United  Gas  Improvement  Co.  60-in.  bar  photometer.  The 
photometric  standards  used  were  10-volt  tungsten  lamps,  pre- 

CANDLEPOWER  OF  VARIOUS  TYPES  OF  PORTABLE  MINERS'  LAMPS 

Candle-        Cost  per 
power       Shift,  Cents 

Miners'  open  oil  cap  lamp 1 . 50  2.4 

Miners'  open  acetylene  cap  lamp 5.00  1.5 

Electric  cap  lamp 1 . 10 

Davy  safety  lamp 0 . 12 

Clanny  safety  lamp 0 . 35 

Wolf -type  safety  lamp 0 . 65 

Akroyd  and  Best  safety  lamp 1 . 10 

T.  M.  Chance  acetylene  safety  lamp 3.80 

NOTE. — The  above  candlepowers  are  in  no  sense  maximum,  but  are  the  average  values 
over  the  field  illuminated  by  the  lamp  in  question  and  have  been  obtained  from  many 
determinations.  These  are  the  value  that  may  be  expected  to  be  realized  in  practice 
under  working  conditiona 


MISCELLANEOUS  INSIDE  COSTS  459 

pared  and  calibrated  by  the  National  Lamp  Works,  and  stand- 
ard sperm  candles. 

No  estimate  of  the  cost  per  day  of  electric  cap  or  flame 
safety  lamps  is  given  in  the  table.  The  labor  charge  on  both 
the  electric  cap  and  flame  safety  lamps  is  so  large  and  varies 
so  much  with  the  size  of  the  installation  that  such  figures  as 
could  be  given  would  have  but  little  meaning. 

Portable  electric  lamps. — The  qualifications  of  portable 
electric  lamps  can  be  grouped  under  three  main  heads  as  fol- 
lows :  Weight,  cost  and  capacity. 

The  weight  of  a  lamp  can  be  easily  ascertained  and  each 
prospective  user  of  a  lamp  must  decide  for  himself  whether 
or  not  its  weight  is  excessive.  Under  the  head  of  cost  would 
be  included  the  first  cost  of  the  equipment,  as  well  as  all 
proper  charges  for  operating  and  maintaining  the  lamp.  Some 
of  these  charges  will  vary  with  each  installation  and  whether 
or  not  the  cost  is  excessive  will  depend  somewhat  upon  the 
conditions  which  surround  each  case. 

The  capacity  of  the  lamp  is  taken  to  mean  its  ability  to 
produce  a  certain  amount  of  light  for  a  definite  number  of 
hours  per  day,  every  day  in  the  year  if  need  be.  A  lamp  that 
can  do  this  with  the  fewest  interruptions  has  the  greatest 
capacity  for  performing  the  duty  for  which  it  is  intended.  The 
capacity  of  a  lamp  as  thus  defined  takes  into  consideration  not 
only  the  ampere-hour  capacity  of  the  battery  and  the  efficiency 
of  the  lamp  bulb,  but  also  the  life  of  battery  plates,  the 
mechanical  strength  of  parts  and  the  resistance  to  wear  and 
tear. 

We  need  to  define  (1)  what  is  the  proper  amount  of  light 
for  a  lamp  to  give;  (2)  the  proper  time  it  should  burn  each 
day,  and  (3)  what  are  reasonable  interruptions  of  service  and 
how  often  they  may  occur. 

Proper  care  of  the  lamps  has  considerable  effect  on  the 
reliability  of  the  service.  One  of  the  large  German  mines, 
having  several  thousand  electric  lamps  in  daily  use,  reports 
that  at  first  about  5  per  cent  of  all  lamps  taken  into  the  mine 
at  the  beginning  of  the  shift  were  returned  at  the  end  of  the 
same  shift,  either  burning  poorly  or  not  at  all.  By  a  careful 
study  of  all  details  in  the  lamp  house  and  by  putting  a  skilled 
man  in  charge  of  the  lamp-house  work,  this  percentage  was 


460  COAL  MINING  COSTS 

reduced  to  less  than  1.5,  with  the  expectation  that  it  would 
eventually  drop  below  1  per  cent. 

While  the  first  cost  of  electric  lamps  is  undoubtedly  higher 
than  that  of  those  burning  benzine,  the  cost  of  operation,  in- 
cluding maintenance,  is  claimed  to  average  from  10  to  15  per 
cent  less.  The  cost  of  the  electrical  energy  is  small  and  the 
cost  of  maintenance  consists  about  one-third  of  labor  and  two- 
thirds  of  renewal  of  parts,  and  depreciation. 

Of  especial  importance  is  the  cost  of  renewing  the  electrodes 
of  the  storage  batteries,  replacement  of  complete  lamps,  which 
are  broken  on  account  of  rough  handling  and  accidents,  and 
renewing  the  incandescent  bulbs. 

The  life  of  the  electrodes  for  lead  cells  ranges  from  about 
100  to  400  shifts,  depending  entirely  upon  the  treatment  which 
they  receive. 

The  replacement  of  complete  lamps  which  are  broken  on 
account  of  rough  handling  and  accidents  undoubtedly  varies 
more  or  less  in  accordance  with  the  character  of  the  work  per- 
formed in  the  mine.  European  practice  shows  that  about  0.1 
per  cent  of  all  lamps  per  shift  are  lost  in  this  manner. 

The  incandescent-lamp  renewal  already  has  been  expressed 
in  figures,  in  connection  with  the  reliability  of  service.  Excel- 
lent results  have  been  obtained,  the  average  life  of  the  lamps 
being  approximately  1000  hours. 

The  Manlite  lamp  complete,  including  the  battery,  cable 
and  head  piece,  weigh  3y2  Ik- ;  °f  this  the  head  piece  weighs 
4  oz.  The  light  of  this  lamp  is  brilliant,  but  soft,  and  at  the 
same  time  absolutely  steady  and  unflickering  for  at  least  12  hr. 
per  single  charge  of  battery.  The  curves  of  discharge  of  the 
latter  shown  in  Fig.  3,  show  how  it  improves  in  service. 

On  the  first  discharge  the  voltage  falls  to  1.8  in  a  little  more 
than  8  hr.  The  second  discharge  reaches  the  same  point  in  a 
little  over  half  as  long  again.  The  progress  is  thereafter  not 
so  rapid,  but  by  the  twelfth  discharge  the  voltage  is  main- 
tained at  over  1.8  for  more  than  153/2  hr.  By  the  thirty-seventh 
discharge  that  period  is  extended  over  the  sixteenth  hour. 

But  these  curves  are  based  on  a  discharge  of  1  amp.,  whereas 
the  lamps  only  use  0.78  to  0.83  amp.,  and  consequently  the 
voltage  is  maintained  for  much  longer  periods  than  those  men- 
tioned. The  curve  of  charge  with  2  amp.  shows  a  slow  increase 


MISCELLANEOUS  INSIDE  COSTS 


461 


in  voltage  till  the  fifth  or  sixth  hour,  when  the  rise  becomes 
quite  rapid  till  the  seventh  hour. 

The  batteries  will  supply  light  after  10  charges  and  dis- 
charges for  20  hr.  at  a  stretch.  This  has  been  shown  by  tests 
duly  authenticated  by  mining  companies  which  have  used  the 
outfits. 

The  bulb  employed,  and  approved  by  the  Bureau  of  Mines, 
has  a  guaranteed  life  averaging  600  burning  hours.  It  is  held 
in  the  burning  position  by  a  perfectly  ground  crystal  lens 
carried  securely  in  place  by  the  lens  holder,  the  whole  being 
sealed  so  that  the  miners  cannot  tamper  with  it. 

Accurate  figures  of  the  actual  cost  of  upkeep  of  the  lamp 
in  continuous  practical  service  for  periods  ranging  from  six 
months  to  over  a  year  at  various  large  mines  have  been  col- 


180    I    I 3  4    5   6 

Time  in  Hours 

FIG.  3. — Curves  showing  how  battery  maintains  its  voltage  for  a  lengthened 
period  after  repeated  use. 

lected  which  show  that  the  cost  of  material  per  lamp-shift 
does  not  exceed  li/2c-  In  a  lamphouse  operating  320  of  these 
lamps  during  five  months  no  battery  repairs  or  renewals  were 
required  aside  from  a  small  quantity  of  electrolyte.  The  cost 
of  other  mechanical  repair  parts  for  the  320  lamps  amounted 
to  only  $15.25. 

The  equipment  is  sold  at  a  reasonable  figure,  and  the  cost 
of  renewal  and  repair  parts  is  equally  low  in  price;  for  in- 
stance, the  cost  of  a  positive  plate  approximates  50c.,  while  a 
set  of  negative  plates  costs  $1. 

At  the  Merchants  Coal  Co.,  Orenda  No.  2  mine,  Boswell, 
Penn.,  250  Edison  lamps  were  installed  on  February  15,  1916, 
at  a  total  cost  of  approximately  $3200.  These  lamps  were  in 
continuous  service  for  about  5  months  during  which  time 
they  were  used  the  equivalent  of  30,450  lamp-shifts.  During 
this  time  the  total  cost  for  bulbs,  cords,  lenses  and  all  other 
repair  parts  was  $126.49,  or  a  trifle  less  than  15c.  per  lamp 


462  COAL  MINING  COSTS 

per  month.  On  the  basis  of  30,450  lamp-shifts,  the  cost  per 
lamp-shift  was  0.4c.  Add  to  this  the  cost  of  lamp  tender, 
current,  approximate  depreciation  and  interest  on  original 
investment,  and  the  total  is  2c.  per  lamp  per  shift. 

This  installation  shows  the  approximate  range  of  cost, 
fixed  charges,  depreciation,  interest  on  original  investment, 
service  and  current. 

The  figures  given  herewith  were  taken  from  the  invoices 
covering  material  shipped  to  the  mine  during  the  period  men- 
tioned, and  no  account  has  been  taken  of  the  material  on  hand 
on  June  1,  1916,  in  either  case;  so  that  there  is  a  satisfactory 
margin,  and  the  costs  shown  are  slightly  in  excess  of  the  actual 
replacements  during  the  period  outlined.  No  expense  has  been 
incurred  for  repairs  or  replacements  of  the  batteries  them- 
selves. 

A  12-months  test  of  electric  lamps  was  made  at  the  Vulcan 
mine,  Newcastle,  Colo.,  of  the  Rocky  Mountain  Fuel  Co.,  about 
1915.  It  is  important  to  note  that  the  installation  was  not 
large,  and  so  the  figures  well  represent  what  might  be  readily 
duplicated  in  a  similar  small  station.  Moreover  it  may  be 
noted  that  the  lamps  were  tended  by  the  regular  lampman  of 
the  mining  company  and  not  by  an  experienced  and  specially 
trained  man. 

The  Vulcan  mine  dips  at  an  angle  of  over  45  deg.,  and  the 
lamps  are  operated  under  the  most  unfavorable  conditions, 
being  subjected  to  the  roughest  usage.  On  December  23,  1914, 
27  lamps  were  put  into  service,  and  monthly  statements  were 
made  out,  showing  not  only  the  number  of  delivered  lamp  shifts 
and  burning  hours  of  the  different  lamps,  but  every  repair 
necessary  and  every  spare  part  consumed. 

The  total  number  of  delivered  lamp-shifts  during  the  12 
months  was  6350,  making  a  total  of  burning  hours  of  61,700. 
The  number  of  lamp-months  was  324,  and  the  cost  of  upkeep 
excluding  labor  was,  as  shown,  $65.52  net  or  20.22c.  per  lamp 
per  month.  This  figure  included  all  material  needed  to  keep  the 
lamps  in  good  working  condition.  Even  lye  solution,  vaseline 
and  acid  received  due  consideration. 

The  average  number  of  lamp-shifts  was  19.6  per  month. 
Thus  the  upkeep  per  shift  per  lamp  was  only  1.03c.  Including 
the  27  bulbs  furnished  with  the  lamps,  113  bulbs  were  rendered 


MISCELLANEOUS  INSIDE  COSTS  463 

valueless  during  the  year.  Of  these  34  were  destroyed  by  the 
miners.  These  may  be  figured  as  being  half-consumed  before 
they  were  destroyed,  and  the  same  assumption  is  made  about 
the  27  bulbs  in  use  at  the  end  of  the  period. 

COST  OF  MATERIAL  AT  SELLING  PRICES  USED  AT  VULCAN  MINE  DURING 
12  MONTHS  ABOUT  1915 

Material  Consumed  and  Destroyed  by  Carelessness : 

86  bulbs,  45c.  each $38! 70 

49  cable  lengths,  4  ft.  each,  20c.  each 9 . 80 

20  spring  terminal  sockets,  12c.  each 2.40 

8  connection  pieces,  20c.  each 1 . 60 

2  lens  holders,  20c.  each .40 

100  lead  seals  for  lamps,  60c.  per  100 60 

15  rubber  corks,  5c.  each .75 

10  lead  check  nuts  with  washer,  13c.  each ....       1 . 30 
17  lamp  holders,  15c.  each 2.55 

2  lenses,  20c.  each 60 

3  lamp  bodies,  70c.  each 2 . 10 

8  celluloid  battery  casings,  $1.40  each 11.20 

43  celluloid  battery  tops,  30c.  each 12.90 

9  oz.  celluloid  strips,  25c.  per  oz 2 . 25 

5i  pt.  celluloid  paste,  $1.75  per  pt 9 . 62 


Gross  total  of  material $96 . 77 

Less  20  per  cent  trade  discount 19 . 35 


Net  total  of  material $77 . 42 

Material  for  Care  of  Batteries: 

10  cans  of  lye  solution,  lOc.  each $1 .00 

8  Ib.  of  vaseline,  8c.  Ib ...          .64 

19  gal.  of  acid,  18c.  gal 3.42 

5.06 


Net  total $82.48 

Material  Destroyed  by  Carelessness  of  Miners: 

34  bulbs,  45c.  each $15.75 

17  lamp  holders,  15c.  each 2 . 55 

3  lenses,  20c.  each 60 

3  lamp  bodies,  70c.  each 2 . 10 

1  lens  holder,  20c.  each 20 


Gross  total  of  material $21 . 20 

Less  20  per  cent  trade  discount 4 . 24 

16.96 


Material  consumed  by  use  during  12  months $65.52 


464  COAL  MINING  COSTS 

Thus  61  bulbs  may  be  figured  as  half-consumed,  which  is 
equivalent  to  about  31  bulbs  wholly  burned  out.  Thus  it  is 
fair  to  assume  that  the  requirements  of  the  year's  running 
were  82  bulbs.  These  delivered,  as  stated,  61,700  burning  hours, 
or  753  hr.  per  bulb,  which  is  a  good  result. 

It  is  interesting  to  note  that  no  battery  plates  had  to  be 
renewed  during  the  whole  year,  and  the  mine  reports  that  all 
the  electrodes  were  still  in  first-class  condition. 

It  will  be  noted  that  82  bulbs  were  sufficient  for  27  lamps, 
so  that  less  than  three  bulbs  served  for  each  lamp  for  one 
whole  year. 

Forty-nine  cable  lengths  had  to  be  renewed  during  the 
trial  year.  Let  it  be  assumed  that  the  27  cables  were  half  worn 
out  when  the  year  was  concluded.  This  will  make  the  equiva- 
lent number  of  cables  a  trifle  under  63.  In  other  words,  as 
the  total  number  of  delivered  lamp  shifts  was  6350,  a  cable 
lasted  for  over  100  shifts,  or  as  19.6  shifts  were  delivered  per 
month,  it  lasted  for  over  five  months,  a  long  life  for  a  part 
under  such  stress. 

This  estimate  does  not  cover  the  maintenance  charges 
though  such  a  small  installation  does  not  give  representative 
results  in  this  respect.  On  the  other  hand  a  new  fastening 
has  been  adopted  since  this  lamp  was  put  in  service  which  will 
greatly  lengthen  the  life  of  these. 

It  has  been  found  that  250  lamps  can  easily  be  tended  by 
one  lampman  at  $75  per  month  and  one  assistant  at  $45  per 
month  or  a  total  of  $120  per  month.  Assuming  that  each  lamp 
is  operated  for  19.6  shifts,  the  250  lamps  deliver  4900  shifts 
for  a  labor  cost  of  $120  or  2.45c.  per  lamp-shift.  Adding  the 
cost  of  upkeep  to  this  the  total  cost  including  all  charges  will 
be  3.48c.  or  roughly  31/2^. 

At  the  Keystone  Coal  and  Coke  Co.'s  Salem  mine,  New 
Alexandria,  Penn.,  200  Edison  electric  safety  mine  lamps  are 
in  service.  One  hundred  of  these  were  installed  September  17, 
1915,  50  were  added  November  18  of  the  same  year  and  an- 
other 50  on  January  6,  1916. 

Including  the  erection  of  charging  racks,  switchboards,  etc., 
the  cost  of  installing  these  200  lamps  was  approximately  $2000. 
From  September  17,  1915,  when  the  first  lot  was  installed, 
until  June  1,  1916,  the  only  expense  for  maintenance  and 


MISCELLANEOUS  INSIDE  COSTS  465 

upkeep,  all  necessary  supplies  and  repair  parts,  such  as  cords, 
bulbs  and  lenses,  was  $383.87,  which  is  only  a  trifle  over  24c. 
per  lamp-month. 

During  the  period  from  September  17,  1915,  to  June  1, 
1916,  there  was  a  total  of  34,250  lamp-shifts,  so  that  the  actual 
cost  per  lamp-shift  for  maintenance  and  upkeep  during  that 
time  was  l%c.  The  total  cost  for  service,  including  salary 
of  lamp  tender,  current,  interest  on  investment  and  deprecia- 
tion, was  2%c.  per  lamp-shift. 

As  these  lamps  were  in  continuous  service  for  a  period  of 
6  mo.,  and  some  for  more  than  8  mo.,  these  figures  are  interest- 
ing, in  view  of  the  fact  that  they  represent  the  maximum  cost 
at  any  time  during  the  operation  of  this  type  of  lamp,  since 
from  its  very  construction,  there  can  be  no  further  deprecia- 
tion or  necessary  repairs  except  those  included  in  the  amount 
given. 

This  same  company  also  had  250  lamps  in  operation  at  its 
Crow's  Nest  mine  in  1915,  the  operating  costs  of  which  are  of 
interest. 

In  a  six-months'  period  the  lamps  were  burned  for  34,419 
lamp-shifts.  The  list  cost  of  the  spare  parts  totals  $521.83. 
A  discount  of  20  per  cent  may  be  figured  on  these  prices,  leav- 
ing the  net  cost  of  parts  and  acid  consumed  $417.45. 

But  of  this  some  material  was  broken  while  in  the  care  of 
the  miners: 

6  lamp  bodies,  at  70c $    4 . 20 

20  lenses,  at  20c 4.00 

218  bulbs,  at  45c 98. 10 

3  safety  devices,  at  25c .75 

2  lamp  holders,  at  15c .30 


$107.35 
Less  20  per  cent  discount $21 . 47 


Net  cost  of  material $85 . 88 

This  leaves  $331.57  of  the  cost  chargeable  against  the  lamp- 
station  expense,  and  dividing  the  sum  thus  obtained  by  the 
number  of  lamp  shifts  34,419,  the  cost  of  material  obtained  is 
0.963c.,  or  less  than  Ic.  per  lamp-shift. 

It  will  be  noted  that  the  figures  given  include  not  merely 
what  are  known  as  spare  parts  and  repairs,  but  sulphuric  acid 


466  COAL  MINING  COSTS 

and  celluloid  paste  also.  Thus  all  possible  material  charges 
are  included,  though  it  will  be  observed  that  the  cost  of  current, 
amortization,  interest  and  labor  of  lampman  are  not  figured. 

It  is  noticeable  that  the  cost  of  electric  bulbs  and  cable 
are  the  two  principal  items  in  the  list.  Schedule  6A  of  the 
United  States  Bureau  of  Mines  requires  that  the  average  life 
of  lamp  bulbs  shall  be  not  less  than  300  hr.  for  primary  and 
acid  storage  batteries  and  not  less  than  200  hr.  for  storage 
batteries  using  an  alkaline  solution.  Not  more  than  5  per  cent 
of  the  bulbs  examined  shall  give  less  than  250  hr.  life  with  acid 
batteries,  nor  less  than  170  hr.  life  with  batteries  having  an 
alkaline  electrolyte. 

If  the  number  of  working  days  in  the  month  is  taken  at 
25  and  the  number  of  shift-hours  at  12,  there  would  be  300  hr. 
during  which  the  lamps  would  be  in  use  in  any  month.  Under 
such  circumstances  a  bulb  of  300-hr,  capacity  would  have  to 
be  renewed  every  montfr,  or  12  times  in  a  year.  If  the  price 
of  each  bulb  is  36c.,  the  cost  per  annum  would  be  $4.32  per 
lamp  per  year.  This  maintenance  would  be  far  too  heavy,  as 
it  covers  the  cost  of  bulbs  only. 

At  Crow's  Nest  the  total  number  of  approved  bulbs  re- 
ceived up  to  September  18,  1915,  was  995.  Some  of  these  of 
course  were  mounted  in  new  lamps  and  others  were  placed  in 
stock.  On  September  18  there  were  157  bulbs  in  reserve. 
Thus  there  were  838  bulbs  broken,  consumed  or  in  use.  The 
miners  had  broken  218  bulbs  in  service.  It  seems  conservative 
to  rate  half  of  these  as  burned-out  bulbs.  The  bulbs  were 
destroyed  by  the  mishandling  of  the  miners,  but  not  being 
new  they  would  not  have  burned  as  long  as  bulbs  from  stock, 
even  if  they  had  not  been  injured.  The  loss  from  negligence 
can  therefore  be  calculated  as  being  equivalent  to  109  bulbs. 

This  can  be  deducted  from  the  loss  as  previously  obtained, 
leaving  729  bulbs  destroyed  by  burning  out.  The  same  assump- 
tion, that  the  bulbs  when  unbroken  have  still  half  their  life 
unexpired,  may  be  applied  to  the  bulbs  in  the  lamps.  There 
are  250  of  these.  It  will  be  assumed  that  their  life  is  equiva 
lent  to  that  of  125  bulbs  fresh  from  stock.  These  bulbs  can 
be  deducted  from  the  number  already  obtained  and  the  bulbs 
actually  consumed  will  be  604. 

The  number   of  lamp-shifts  of  these   604  bulbs  is  34,419. 


MISCELLANEOUS  INSIDE  COSTS 


467 


Each  shift  is  taken  at  11  hr.,  giving  378,609  lamp  hours,  or 
627  burning  hours  per  bulb,  or  57  shifts  of  11  hr.,  instead  of 
the  minimum  average  as  required  by  the  bureau  for  an  acid 
battery,  namely,  300  hr. 

Assuming  the  number  of  shifts  in  a  year  to  be  300,  then 
5.26  bulbs  will  be  needed  per  year,  which  at  36c.  would  entail 
an  expenditure  of  $1.89.  Where  the  Mannesmann  Light  Co. 
maintains  its  lamps  at  so  much  per  shift  it  stipulates  that  the 
operating  corporation  shall  charge  its  miners  with  the  net 
amount  of  all  material  destroyed  through  the  carelessness  of 
these  employees.  It  is  fully  justified  in  assuming  therefore 
that  the  estimates  of  lamps  broken  by  miners  is  not  in  any 
way  excessive. 

The  cable  used  is  marked  as  5  ft.  long.  As  a  matter  of 
fact  it  need  only  be  4%  ft.  The  additional  6  in.  is  added  for 
conservative  estimation.  The  cable  is  the  one  part  of  the 
lamp  still  furnishing  a  small  problem  to  the  manufacturers, 
and  efforts  are  being  made  to  solve  it  in  a  more  satisfactory 
degree. 

MATERIAL  DESTROYED  BY  MINERS  AT  CROW'S  NEST  MINE  FROM  MARCH  22 
TO  SEPTEMBER  18,  1915 


Period 

Lamp 
Bodies 

Lenses 

Bulbs 

Safety 
Devices 

Lamp 
Holders 

March  22  to  April  30  
During  May  

1 
3 

5 
4 

50 
52 

1 
1 

1 

During  June  

3 

45 

During  July  
During  August 

1 

1 

6 

34 
23 

1 

•• 

September  1  to  September  18 

1 

1 

14 

1 

6 

20 

218 

3 

2 

According  to  observations  made,  a  lamp  cable  is  bent  on 
the  miner 's  back  about  7000  times  during  the  length  of  a  single 
shift.  It  is  easy  to  understand,  therefore,  to  what  a  severe 
service  it  is  subjected.  It  is  not  so  much  the  kind  and  quality 
of  the  cable  which  is  in  question.  The  important  matter  is  the 
manner  in  which  the  cable  is  attached  to  the  lamp  and  battery 
casing.  The  tests  mentioned  are  being  made  from  this  point 
of  view. 


468 


COAL  MINING  COSTS 


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MISCELLANEOUS  INSIDE  COSTS  469 

It  is  necessary  to  consider  not  only  cost  of  material  but 
the  labor  and  amortization  charges.  The  experience  of  the 
Mannesmann  Light  Co.  has  been  that  one  man  can  handle 
at  least  150  electric  safety  cap  lamps  and  give  perfect  satis- 
faction. Thus  300  lamps  could  be  tended  by  one  lampman 
and  his  assistant.  If  these  men  are  paid  $3  and  $2  a  day 
respectively,  and  25  days  are  figured  to  a  month,  the  monthly 
payroll  will  be  $125.  If  they  work  to  capacity — that  is,  charge 
and  clean  300  lamps  for  each  of  the  25  shifts — each  lamp- 
shift  labor  charge  will  be  1.66c.,  or  about  l%c. 

In  figuring  the  battery  maintenance  charges  no  considera- 
tion has  been  given  to  battery  renewals,  because  in  the  six 
months  during  which  the  lamps  were  installed  at  the  Crow's 
Nest  mine  the  batteries  did  not  need  any  repairs.  But  the 
time  will  come  when  they  will,  so  it  is  well  to  be  safe  and  to 
double  the  repair  bill  arbitrarily,  making  that  item  2c.  Add- 
ing the  labor  charge  to  this,  the  cost  of  operating  is  3.7c.  per 
lamp-shift. 

The  custom  in  most  American  coal  mines  is  to  charge  the 
miner  at  least  5c.  for  the  use  of  a  safety  lamp  during  one  shift. 
Taking  5c.  as  the  charge,  the  balance  left  for  the  operator 
would  be  1.3c.,  and  this  would  have  to  cover  amortization  and 
interest  on  invested  capital.  A  small  lamp  plant  of  300  lamps 
running  250  shifts  per  year  would  give  a  total  of  90,000  lamp- 
shifts,  which  at  1.3c.  would  provide  $975  for  amortization  and 
interest. 

Such  a  plant  for  the  Manlite  lamps,  including  the  lights 
themselves,  would  cost  about  $2750.  An  amortization  charge 
of  30  per  cent  and  6  per  cent  interest  would  amount  to  $962.50. 
So  it  will  readily  be  seen  that  the  electric  cap  lamp  not  only 
gives  a  good  light  and  assures  safety  from  gas  and  other 
causes,  but  also  is  at  5c.  per  lamp-shift  by  no  means  an  unsatis- 
factory investment. 

At  the  mine  of  the  Federal  Coal  and  Coke  Co.,  Grant  Town, 
W.  Va.,  an  installation  was  started  in  October,  1914,  with  80 
electric  hand  lamps  and  this  number  was  increased  to  560 
lamps  in  April,  1915.  A  record  is  therefore  available  for  this 
considerable  number  of  lamps,  for  an  entire  year. 

The  total  cost  of  material  shipped  to  this  particular  mine 
during  the  12  mo.  in  question  amounted  to  $1769.51.  During 


470  COAL  MINING  COSTS 

this  period,  the  sum  charged  to  the  miners  on  the  pay  roll, 
for  material  intentionally  or  carelessly  destroyed,  amounted 
to  $196.88  so  that  the  total  cost  of  material  to  the  company, 
ranging  over  a  period  of  a  full  year,  was  $1572.63.  In  this 
amount  no  consideration  was  taken  of  the  materials  and  repair 
parts  still  on  hand  in  the  stock-room  of  the  mine,  at  the  expira- 
tion of  this  period.  The  value  of  these  materials  and  parts 
was  quite  appreciable. 

During  this  period  the  number  of  lamp-shifts  furnished  to 
the  mine  amounted  to  141,316,  so  that  the  cost  per  lamp  shift 
for  maintenance  and  upkeep  would  amount  to  l.llc. 

The  above  figures  do  not  include  labor.  The  lamps  at  the 
above-named  mine  were  in  charge  of  an  expert  lamp  man  with 
from  2  to  3  assistants,  according  to  the  conditions  prevailing 
at  the  mine,  but  the  total  wages  paid  out  at  the  lamp  house 
at  no  time  amounted  to  more  than  about  $200  per  month.  This 
would  amount  to  approximately  1.4c.  per  lamp-shift. 

The  total  cost  of  operation  for  material  and  labor  there- 
fore amounted  to  approximately  2y2c.  per  lamp-shift. 

These  lamps  have  been  used  daily  without  interruption 
since  they  were  first  installed.  It  should  also  be  noted  that 
only  a  small  percentage  of  the  material  actually  used  was 
charged  to  the  miners;  viz.,  about  11  per  cent. 

Oil  and  acetylene  lamps. — The  Union  Pacific  Coal  Co.,  pre- 
pared the  following  figures  as  to  the  cost  of  burning  acetylene 
and  oil  lamps  in  1914.  No  attempt  has  been  made  to  compare 
the  candlepower-hours.  Such  a  comparison  would  be  more 
favorable  to  the  acetylene  lamp  than  is  here  given.  The  period 
on  which  the  estimates  are  based  is  a  day  of  8  hr. 

COST  OF  ACETYLENE  AND  OIL  LIGHTING 


Name  of  Mine 

Carbide  Lamp 

Lard-oil  Lamp 

Reliance                          

$0  015 

$0  048 

Rock  Springs 

0  055 

0  075 

Cumberland  
Superior 

0.035 
0  034 

0.045 
0  075 

Hanna                         

0  034 

0  070 

Averaere.  . 

0.0346 

0.0626 

MISCELLANEOUS  INSIDE  COSTS  471 

The  estimates  are  based  on  a  charge  of  8Y3c.  per  Ib.  for 
carbide  and  65c.  to  75e.  per  gal.  for  lard  oil. 

The  oil  lamp  is  rapidly  disappearing  from  the  mines  in  this 
country  but  for  the  benefit  of  those  mines  still  using  the  oil 
lamp  the  following  particulars  are  given. 

Oil  lamps. — The  viscosity  of  an  oil  is  the  property  of  the 
different  particles  to  hold  together  and  at  the  same  time  adhere 
to  other  bodies.  It  may  be  called  the  "fluidity"  or  "thick- 
ness" of  an  oil.  Oil  of  high  viscosity  does  not  flow  so  freely 
as  an  oil  of  low  viscosity.  Viscosity  has  been  the  cause  of 
complaint  with  miners'  oil,  otherwise  satisfactory,  as  for  in- 
stance, an  oil  which  gave  the  following  results  upon  test: 
Viscosity,  16.2;  density,  22  deg.  Be.;  grams  oil  burned  per 
minute,  .2;  height  of  flame,  4  inches.  This  oil  was  the  cause 
of  dissatisfaction  at  several  of  the  mines  the  complaint  being 
that  it  did  not  keep  lighted  on  headings  and  air-courses  where 
strong  air-currents  were  met  with,  that  it  was  hard  to  relight, 
and  that  the  oil  did  not  feed  up  the  wick  fast  enough.  While 
the  flash  point  of  this  oil  was  somewhat  to  blame  for  the  trouble, 
it  was  principally  due  to  the  high  viscosity. 

The  congealing  point  of  oil  is  an  important  factor  in  the 
winter  months.  It  is  the  temperature  at  which  the  oil  will 
solidify,  and  if  this  point  is  too  high  it  is  the  cause  of  consider- 
able trouble  both  to  the  miner  and  storekeeper,  which  results 
in  a  loss  of  time  and  waste  of  oil.  Although  the  temperature 
inside  the  mines  varies  but  little,  there  are  many  cases  where 
roadmen,  drivers,  motormen,  and  others,  come  in  contact  with 
air  that  is  almost  as  cold  as  that  outside,  consequently,  when 
buying  winter  oil  there  should  be  a  guarantee  that  it  will  not 
congeal  at  a  temperature  above  32  deg.  F.  and  it  is  advanta- 
geous to  have  this  point  even  lower.  The  winter  oil  used  by  the 
Fairmont  Coal  Co.  shows  a  congealing  point  of  21  deg.  F.  on 
an  average  of  22  samples  tested.  This  allows  it  to  be  stored 
and  handled  in  cold  places  without  freezing. 

The  flash  point  is  the  temperature  in  degrees  Fahrenheit  at 
which  the  oil  will  flash  or  ignite.  For  instance,  an  oil  with  a 
flash  point  of  300  deg.  is  one  that  will  ignite  when  heated  to 
that  temperature,  if  a  flame  be  applied  to  it.  The  flash  point 
decreases  as  the  density  becomes  less;  the  lighter  oils  flashing 
at  lower  temperature.  Some  companies  have  300  deg.  to  425 


472  COAL  MINING  COSTS 

deg.  F.  established  as  the  limit  in  their  specifications  for  miners' 
oil,  that  is,  the  flash  point  must  not  drop  below  300  deg.  nor 
exceed  425  deg.  F. ;  the  lower  limit  being  applied  as  a  measure 
of  safety ;  the  higher  one  as  a  limit  at  which  the  oil  will  readily 
burn  and  relight  quickly. 

The  quantity  of  oil  burned  for  a  given  length  of  time  is 
important  from  the  miner's  point  of  view  (although  it  is  hard 
to  make  him  realize  it).  In  trying  to  comply  with  the  state 
law,  and  also  furnish  an  oil  that  will  give  the  miner  the  best 
results  per  unit  of  cost,  it  is  necessary  to  take  into  considera- 
tion its  efficiency.  The  grams  of  oil  burned  per  minute  depend 
on  the  size  of  the  lamp  used,  the  size  of  the  wick,  and  the 
length  of  the  wick  beyond  the  spout,  as  well  as  the  quality  of 
oil.  With  the  first-mentioned  conditions  remaining  constant, 
the  variation  in  the  grams  burned  per  minute  depends  to  some 
extent  upon  the  density.  With  a  cottonseed  oil  of  high  specific 
gravity  the  oil  burned  per  minute  in  a  standard  No.  2  mine  lamp 
is  about  0.2  gram,  while  in  an  oil  highly  adulterated  with  kero- 
sene, or  a  light,  highly  volatile  oil,  and  of  a  low  specific  gravity, 
the  figure  is  about  0.5  gram. 

The  candlepower  of  oil  naturally  bears  a  close  relation  to 
the  quantity  of  oil  burned,  and  to  get  a  comparative  figure  on 
different  oils  as  to  this  property,  the  candlepower  is  calculated 
to  candlepower  per  gram  per  minute.  This  takes  into  con- 
sideration the  effect  of  different  rates  of  burning.  In  making 
the  test  of  the  candlepower,  standard  photometric  candles  (12 
to  a  pound)  are  used,  in  an  ordinary  photometer,  and  the 
results  check  very  closely.  The  present  specifications  call  for 
an  equivalent  candlepower  per  gram  per  minute  at  8. 

The  height  of  flame  is  also  taken  into  consideration  and 
is  specified  in  purchasing  oil  to  be  between  4  and  5.5  in.  with 
the  wick  .extending  14  in.  beyond  the  end  of  the  lamp  spout, 
with  12-strand  wick.  This  height  has  been  found  to  produce 
the  best  candlepower  with  the  least  amount  of  smoke,  and  if 
the  height  of  flame  is  limited,  the  evolution  of  smoke  is  to 
some  extent  controlled. 

In  testing  miners'  oil  there  is  probably  nothing  of  greater 
importance  than  the  quantity  of  smoke  it  gives  off  while  burn- 
ing. This  smoke  is,  in  a  measure,  indicative  of  the  quality  of 
the  oil,  as  an  inferior  oil  always  makes  denser  smoke  and  a 


MISCELLANEOUS  INSIDE  COSTS  473 

greater  quantity  of  it,  and  it  has  often  been  noted  that  the 
smoke  increases  as  the  oil  deteriorates  in  quality.  It  is  this 
point  which  guides  the  mine  inspector  in  his  inspection  of  the 
oil  used  at  different  mines.  If  the  oil  gives  off  an  excessive 
amount  of  smoke,  he  becomes  suspicious  and  forwards  a  sample 
to  the  head  of  the  department  for  test,  and  it  usually  develops 
that  the  oil  is  adulterated  and  not  up  to  the  standard  required. 
Aside  from  the  dust  and  the  smoke  of  the  powder,  smoke  from 
open  lights  is  the  principal  cause  of  air  vitiation  in  the  work- 
ing places  of  a  mine.  This  is  especially  noticeable  in  places 
where  the  velocity  of  the  current  is  very  low,  and  with  oil  of 
low  grade.  An  oil  to  which  kerosene  has  been  added  (which 
is  common  practice  in  most  regions)  is  the  greatest  offender  in 
this  direction. 

Although  it  is  difficult  to  measure  the  quantity  of  smoke 
given  off  by  any  particular  lamp,  the  following  method,  which 
is  similar  to  the  one  used  by  the  state  mine  inspector,  gives 
comparative  results  and  is  sufficient  for  testing  miners'  oil. 
This  test  is  made  by  observation,  using  pure  cottonseed  oil  as 
the  standard  of  comparison.  The  size  of  lamp,  size  and  length 
of  wick,  and  the  quantity  of  oil  are  the  same  in  each  case, 
and  length  of  the  smoke  pencil  on  the  standard  oil  is  taken 
as  one,  and  the  length  of  the  smoke  pencil  on  the  oil  tested  is 
relatively  greater  or  less.  The  lamps  are  placed  in  a  box  with 
a  glass  front,  and  a  white  cloth,  graduated  in  inches,  serves 
as  a  back-ground.  The  limit  placed  on  the  smoke  is  a  pencil 
about  1.5  greater  than  that  of  the  standard  cottonseed  oil. 
This  test,  however,  is  unnecessary  quite  often  when  other 
requirements  above  mentioned  are  complied  with,  as  they 
govern  to  a  very  great  extent,  the  evolution  of  smoke. 

Apparently  the  most  difficult  requirement  to  meet  is  the 
density,  which  must  not  exceed  24  deg.  Be.  scale,  which  cor- 
responds to  .913  specific  gravity.  Of  many  samples  of  miners' 
oil  of  different  kinds  submitted  by  many  companies,  there  has 
been  only  a  small  percentage  to  come  within  the  limit.  Its 
adoption  as  part  of  the  specifications  of  some  companies  has 
been  due  solely  to  the  fact  that  the  State  Oil  Regulations 
require  it.  Miners'  oil  of  excellent  quality  can  be  obtained  at 
very  reasonable  prices,  only  a  degree  over  the  state  law  limit, 
but  is  not  purchased  because  it  exceeds  24  deg.  Be.  This  test 


474  COAL  MINING  COSTS 

is  made  at  60  deg.  F.,  and  while  certain  properties  do  bear 
some  relation  to  the  density,  there  is  no  room,  apparently, 
why  this  limit  should  have  been  placed  as  low  as  24  deg.  Be. 
Very  seldom  has  a  miners'  oil  sample  been  received  in  the 
laboratory  that  does  not  exceed  this  point.  We  have  found 
oils  up  to  30  deg.  Be.  that  were  highly  satisfactory  for  use  in 
miners'  lamps,  being  pure  unadulterated  oils  "as  free  from 
the  evolution  of  smoke  as  a  standard  cottonseed  oil." 

A  trial  test  was  made  extending  over  several  months  on 
standard  cottonseed  oil,  and  the  reports  from  41  mines  using 
it  were,  in  effect,  that  it  was  of  such  high  density  (low  Baume) 
and  high  flash  point  as  to  be  very  unsatisfactory.  The  flame 
would  not  hold  to  the  wick,  the  lamp  once  out  was  hard  to 
relight,  the  light  furnished  was  of  low  candlepower,  and  the 
cost  was  excessive. 

The  specifications  for  miners'  oil  as  adopted  by  the  Fair- 
mont Coal  Co.  in  1910,  were  as  follows:  Density,  24  deg.  Be. 
or  under;  congealing  point  24  deg.  F.  or  under;  flash  point, 
300  deg.  to  425  deg.  F. ;  flame  height,  between  4  and  5%  in. 
in  a  No.  2  miners'  Star  pattern  lamp  with  ^-in.  wick  projec- 
tion and  a  12-strand  cotton  wick;  equivalent  candlepower  per 
gram  per  minute,  8  or  over ;  smoke,  to  be  light  in  quantity. 

When  a  new  oil  is  to  be  purchased,  samples  of  it  are  re- 
ceived and  tested  and  unless  it  comes  up  to  these  specifications 
it  is  not  considered  by  the  purchasing  agent.  If  a  purchase  is 
made  of  an  oil  that  has  been  found  to  be  all  right,  each  con- 
signment is  tested  as  it  is  received  at  the  supply  department. 
If  it  should  here  fail  to  come  up  to  the  requirements,  as  stated, 
it  is  rejected.  It  is  a  rule  to  test  the  oil  at  every  mine  at 
regular  intervals  as  well  as  each  shipment  received  by  the 
supply  department. 

An  example  of  the  manner  in  which  a  law,  that  was  honestly 
intended  to  be  beneficial  to  the  miner,  failed  of  its  purpose  is 
found  in  the  Pennsylvania  Bituminous  Mining  Law  of  1911. 
This  law  stipulated  that  oils  sold  for  use  in  miners'  lamps 
should  not  yield  more  than  0.11  per  cent  of  soot  when  burned 
in  a  miners'  lamp  under  standard  conditions.  One  of  these 
conditions  was  that  the  flame  of  the  lamp  should  be  l1/^  in. 
long.  Now,  low-grade  oils  when  burned  under  these  conditions 
yield  as  much  as  1  per  cent  of  soot,  while  high-grade  oils  will 


MISCELLANEOUS  INSIDE  COSTS  475 

give  as  little  as  0.03  per  cent.  Thus  it  would  seem,  at  first 
glance,  that  this  law  would  considerably  better  conditions  in 
the  mines. 

Such  is  not  the  case,  however.  Oils  to  pass  this  test  must 
be  very  largely  composed  of  costly  fatty  oils  and  this  so  greatly 
increased  the  cost  to  the  miner  that  he  was  obliged  to  look 
for  some  cheaper  illuminant.  Moreover,  instead  of  a  flame 
iy2  in.  long,  the  miner  burns  one  of  a  maximum  length  be- 
cause he  wants  as  much  light  as  he  can  get.  It  has  been  found 
that  while  costly  oils,  containing  high  percentages  of  fatty 
ingredients,  will  produce  much  less  soot  than  oils  of  medium 
price,  and  less  fatty  material,  when  burned  under  legal  test 
conditions,  these  differences  very  largely  disappear  when  these 
oils  are  burned  under  the  conditions  that  obtain  in  the  mines. 
With  very  long  flames  the  high-priced  oils  still  show  a 
superiority  to  the  medium  grade,  but  the  differential  is  so 
slight  as  to  be  of  little  real  moment. 

Indeed,  the  soot-forming  propensities  of  both  these  oils 
under  the  conditions  of  use  are  so  great  that  it  is  idle  to 
attempt  to  classify  one  as  better  than  the  other.  They  are 
both  very  bad.  Thus  with  a  legal  requirement  of  0.11  per  cent 
soot  or  lower,  we  find  'the  oil  passing  this  test  will  give  about 
8  per  cent  of  soot  when  burned  as  it  would  be  in  the  mines — 
that  is,  with  a  flame  5  to  6  in.  long — while  the  oil  that  will  not 
pass  the  legal  requirement,  giving  under  test  conditions,  let 
us  assume,  0.5  per  cent  soot,  will  make  under  actual  working 
conditions  about  9  to  10  per  cent  soot. 

Cost  of  underground  stables. — The  underground  stable 
shown  in  the  accompanying  illustration,  Fig.  4,  is  designed  to 
meet  the  needs  of  mine  owners  who  wish  to  construct  a  stable 
which  is  thoroughly  modern,  conforming  to  all  the  requirements 
of  the  mine  law,  and  one  which  at  the  same  time  will  not  be 
too  expensive  to  permit  of  its  being  built. 

The  drainage  from  each  stall  is  obtained  by  a  grade  of 
2  in.  in  10  ft.  from  the  manger  to  a  gutter  running  along  the 
back  of  the  stalls.  (Dr.  Charles  A.  Leuder,  professor  of 
veterinary  science  at  West  Virginia  University,  states  as  his 
opinion  that  a  slope  of  from  3  to  4  in.  in  10  ft.  will  not  be 
harmful  to  a  horse  or  mule  standing  in  a  stall  over  night.) 
The  gutter  should  have  a  slope  of  at  least  1%  per  cent.  This 


476 


COAL  MINING  COSTS 


MISCELLANEOUS  INSIDE  COSTS  477 

can  usually  be  secured  naturally  by  the  location  of  the  stable. 
If  this  grade  cannot  be  obtained  easily,  the  gutter  may  be 
made  deep  enough  at  one  end  to  afford  the  proper  slope.  In 
this  case  all  or  a  part  of  it  should  be  covered.  The  gutter 
leads  to  a  sump  from  which  the  water  is  pumped  or  drained 
naturally  to  the  main  sump. 

At  the  entrance  to  the  stable  is  a  washroom  with  concrete 
floor  and  walls,  the  latter  being  from  4  to  5  ft.  high.  This  may 
be  made  to  drain  either  to  the  center  or  to  one  corner  and  from 
thence  to  the  sump.  The  horses  or  mules  are  thoroughly 
washed  here  by  means  of  a  hose  while  they  are  being' watered. 
Plenty  of  clean  water  for  this  purpose  is  usually  available 
from  water  rings  around  the  shaft.  If  it  is  deemed  expedient, 
a  small  reservoir  in  connection  with  the  water  rings  may  be 
constructed  at  a  point  sufficiently  high  to  secure  plenty  of  pres- 
sure. A  water  pipe  runs  the  full  length  of  the  stable,  with 
taps  for  connecting  a  hose  at  points  convenient  for  washing 
out  the  stalls.  An  additional  watering  trough  is  placed  opposite 
the  washroom  for  convenience  in  watering  the  animals  as  they 
are  taken  out  in  the  morning. 

The  stalls  are  5  ft.  6  in.  wide  and  7  ft.  8  in.  long  exclusive 
of  the  mangers.  This  gives  space  ample  for  the  comfort  of 
even  the  largest  size  of  horse  used  in  the  mines.  From  the 
back  of  the  stalls  to  the  track  the  distance  is  2  ft.  3  in.  Be- 
tween the  track  and  the  wall  there  is  2  ft.  7  in.  of  clearance. 
This  allows  plenty  of  space  for  a  large  feed  truck. 

The  side  walls  are  of  concrete  and  are  18  in.  thick.  This 
gives  a  good  substantial  wall  which  is  easily  built.  The  roof 
is  an  elliptical  four-ring  brick  arch,  this  type  being  recom- 
mended because  it  gives  the  maximum  of  useful  space  in  pro- 
portion to  the  excavation  required.  Brick  is  used  because  it 
is  believed  that  an  arch  may  be  more  readily  constructed 
underground  of  this  material  than  of  concrete. 

The  floor  is  of  paving  brick  carefully  laid  with  cement 
mortar  on  a  6-in.  concrete  base.  This  is  used  instead  of  con- 
crete because  it  is  believed  that  it  wears  better  and  gives 
a  better  surface  for  the  animals  to  stand  on.  Dr.  Charles  A. 
Leuder,  who  has  been  referred  to  before,  states  that  the  hard- 
ness of  a  brick  or  concrete  floor  will  not  be  harmful  to  a  horse 
or  mule.  One  of  the  best-equipped  shaft  mines  in  the  Con- 


478  COAL  MINING  COSTS 

nelsville  field  in  Pennsylvania  has  an  underground  stable 
iioored  with  brick.  Another  up-to-date  plant  in  the  Fairmont 
region  in  West  Virginia  has  a  stable  with  a  concrete  floor.  If 
concrete  is  desired,  it  will  be  a  simple  matter  in  this  design 
to  leave  out  the  brick  and  put  the  proper  surface  on  the 
concrete. 

The  mangers  and  feed  boxes  are  simply  and  substantially 
constructed  of  l^-in.  gas  pipe  27  X  32-in.  sheet-iron  plates  and 
2-in.  mesh  wire  screen,  securely  held  in  place  by  means  of  straps 
embedded  in  the  concrete  wall  and  sub-floor.  Standard  pipe 
fittings  are  used,  and  no  difficulty  should  be  encountered  in 
constructing  these  mangers. 

Ventilation  is  secured  by  means  of  a  box  regulator  placed 
at  the  back  of  the  stable.  This  should  be  so  arranged  that  it 
can  be  conveniently  closed  when  the  animals  are  first  brought 
into  the  stable  in  the  evening,  in  order  that  drafts  may  be 
kept  away  from  them  while  they  are  cooling  off.  This  should 
be  reopened  in  about  half  an  hour  and  allowed  to  remain  open 
all  night.  The  stable  is  located  so  as  to  secure  a  separate  split 
of  air  from  the  intake.  This  can  usually  be  arranged  with- 
out difficulty.  Owing  to  the  fact  that  the  partitions  between 
the  stalls  consist  of  gas  pipes  suspended  from  the  roof  and 
mangers,  the  entire  stable  is  open  to  a  free  sweep  of  air,  which 
insures  fresh  air  for  the  animals  even  when  lying  down. 

This  form  of  partition  is  in  use  at  the  Continental  No.  1 
mine  of  the  H.  C.  Frick  Coke  Co.  The  stable  boss  at  this  mine 
states  that  there  has  never  been  any  trouble  owing  to  the 
horses  kicking  each  other  and  that  the  partitions  are  entirely 
satisfactory.  If  any  difficulty  is  anticipated,  it  is  suggested 
that  the  partitions  be  made  more  secure  by  suspending  from 
the  gas  pipe  a  wire  screen  similar  to  that  used  in  the  construc- 
tion of  the  mangers.  This,  when  hung,  will  resemble  somewhat 
the  danger  signals  commonly  used  by  the  railroads  to  warn 
trainmen  of  the  nearness  of  an  overhead  bridge  or  tunnel. 

Owing  to  the  wide  variation  in  different  localities  of  the 
cost  of  labor  and  materials,  no  attempt  has  been  made  to  esti- 
mate accurately  the  cost  of  the  stable.  An  approximate  esti- 
mate of  the  cost  of  such  a  stable  with  35  stalls,  computed  on 
what  is  taken  as  an  average  of  labor  conditions  in  West 
Virginia  in  1916,  is  $3800.  This,  of  course,  does  not  include 


MISCELLANEOUS  INSIDE  COSTS 


479 


excavation,  as  it  would  be  impossible  to  estimate  this  without 
a  knowledge  of  the  thickness  of  the  seam  and  other  conditions. 

The  stable  drawings  shown  in  Figs.  5  to  10  were  adopted  as 
standard  by  one  bituminous  and  two  anthracite  coal-mining 
companies  about  1912.  The  design  is  such  as  to  meet  the  re- 
quirements of  the  mine  inspector,  mine  manager  and  veter- 
inary surgeon.  The  stable  is  of  fireproof  material  through- 
out except  the  top  floor  of  the  stalls,  which  for  the  comfort 
of  the  mules  is  made  of  plank. 

Fig.  5  is  a  ground  plan  of  the  stable.  It  is  designed  on 
the  unit  plan,  so  that  the  drawing  is  available  for  a  stable 


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Plan 
FIG.  5. — Plan  of  two  stalls  of  another  type  of  underground  stable. 

to  contain  one  or  any  number  of  mules.  Ample  room  is  pro- 
vided in  front  and  behind  the  stalls.  It  will  be  noticed  that 
a  good  concrete  floor  is  first  laid,  in  which  are  imbedded 
the  tracks  for  the  feed  cars.  Facing  and  also  at  the  back 
of  the  stalls  are  concrete  gutters  for  carrying  out  the  water 
daily  used  to  cleanse  the  stable. 

Each  stall  is  8  ft.  4  in.  long  and  5  ft.  wide,  and  so  is 
commodious  enough  for  the  largest  of  mine  mules.  The  con- 
crete piers  for  the  center  posts  are  clearly  shown  in  the 
drawing. 

Fig.  6  is  a  cross-section  of  the  stable.  The  thickness  of 
the  concrete  floor,  position  of  drain  pipe,  slight  pitch  of  the 


480 


COAL  MINING  COSTS 


plank  floor,  car  tracks,  passage  ways,  gutters,  manger,  feed 
box,  center  posts,  end  arches,  and  the  2-in.  gas  pipe  which 
alone  forms  the  partition  between  the  stalls,  are  all  clearly 
outlined  in  the  drawing. 

The  single  gas  pipe  is  an  improvement  over  the  old  style 
high  board  partition,  as  it  offers  no  resistance  to  the  free 
circulation  of  air,  and  the  building  is  always  free  from 
obnoxious  odors.  When  it  is  necessary  to  stable  a  fractious 
mule,  he  is  usually  placed  in  one  of  the  end  stalls,  and  the 


Cross  Section  on  Center  Line  of  Stall 

FIG.  6. — Cross  section  of  stable  shown  in  Fig.  5  showing  manger  and  drainage 

system. 

gas-pipe  partition  is  reinforced  by  a  strap-iron  lattice-work, 
which  effectually  prevents  him  from  annoying  his  neighbors. 

Fig.  7  is  a  detail  drawing  showing  side  view  and  plan  of 
the  center  post.  This  exemplifies  a  method  of  supporting 
a  stable  roof  which  is  both  effective  and  inexpensive,  and 
in  the  eight  stables  where  it  has  been  tried,  no  failures  have 
been  recorded.  It  may  here  be  noted  that  the  end  arches 
extend  throughout  the  entire  length  of  the  stable,  serving 
the  double  purpose  of  sealing  any  coal  measures  and  sup- 
porting the  mine  roof. 

Fig.  8  gives  the  construction  details  in  a  longitudinal  sec- 
tion through  the  center  posts.  The  method  of  supporting  the 


MISCELLANEOUS  INSIDE  COSTS 


481 


T-rail  laggings,  and  the  position  of  the  mangers  are  clearly 
shown.  Fig.  9  is  a  plan  of  the  arches  with  the  T-rails  in 
position. 


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FIG.  8. — End  elevation  of  arches  for  stable  shown  in  Figs.  5  and  6. 

Fig.  10  supplies  all  needed  details  of  construction  of  the 
manger  with  its  feed  box,  which  have  been  designed  with  a 


482 


COAL  MINING  COSTS 


view  to  sanitation  as  a  leading  desideratum.    They  are  usually 
constructed  in  the  shops  and  sent  to  the  mines  with  all  parts 


FIG.  9. — Plan  of  arches  in  Fig.  8  showing  T-rail  reinforcing. 

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End  Elevation 
FIG.  10. — Detail  of  manger  and  feed  box  for  stable  shown  in  Figs.  5  and  6. 

marked    to    facilitate    in    erecting.      The    work    of    installing 
them  is  thus  simplified,  and  can  be  done  by  the  average  mine 


MISCELLANEOUS  INSIDE  COSTS  483 

timberman.  Particular  attention  is  called  to  the  hole  provided 
in  the  bottom,  affording  a  convenience  for  cleaning  same. 

When  construction  of  the  building  is  completed,  the  sheet 
iron  which  forms  the  temporary  support  for  the  concrete 
arches  is  not  withdrawn,  but  is  left  in  position  and  when  white- 
washed serves  as  an.  efficient  reflector  for  the  electric  lights 
which  are  provided  for  each  stall.  A  32-cp.  lamp  with  a  heavy 
guard  is  provided  with  a  No.  14  (BX)  extension  cord  10  ft. 
long.  With  this  the  stableman  can  make  a  careful  inspection 
of  the  mule  as  it  returns  from  its  day's  toil. 

The  following  is  the  bill  of  material  for  one  stall  only, 
and  an  equal  increase  should  be  made  for  each  additional  mule 
for  which  accommodation  is  required. 

BILL  OF  MATERIAL  FOR  ONE  STALL 
40  bags  of  cement.     5  tons  of  sand. 
14 — T-rail  laggings — 4  ft.  6  in.  long  (40-lb.  rail) 
70  ft. — odd  lengths  old  T-rail  (for  reinforcing). 
1 — 10  ft.  length,  8  in.  dia.,  cast-iron  column  pipe 
13  pieces  of  sheet  iron — 4  ft.  6  in.  long  XI  ft.  6  in.  wide  (for  arches) 

2  pieces  of  No.  8  sheet  iron  5  ft.  long X 3  ft.  wide. 

1  piece  of  No.  10  sheet  iron  2  ft.X2  ft.  for  feed  box. 
5  floor  planks  8  ft.  4  in.Xl  ft.X4  in.  thick. 
1  bolt — 1  ft.  83^  in.  longXf  in.  diam. 
1  bolt — 1  ft.  5^  in.Xf  in.  diam. 

1  bolt — 1  ft.  2^  in.  longXf  in.  diam. 

3  bolts— 1  ft.  3f  in.  longXf  in.  diam. 
3  bolts — 4  in.  longXf  in.  diam. 

2  bolts — 3  in.  longXf  in.  diam. 
28  rivets  1  in.X^  in.  diam. 

1  piece  2  in.  diam.  gas  pipe,  7  ft.  8  in.  long  (for  partition  between  stalls). 

1  piece  3  in.  diam.  gas  pipe,  5  ft.  9|  in.  long,  (for  drain). 

2  iron  straps,  2  ft.  10  in.  long X 2  in.  wide X  \  in.  thick. 
2  iron  straps,  2  ft.  9 \  in.  long X 2  in.  wideX^  in.  thick. 
10  lin.  ft.  of  strap  iron  2  in.  wideXs  in.  thick. 

1  piece  1-in.  mesh  segment  for  tray  4  ft.  longXl  ft.  2  in.  wide. 

2  brackets. 

Overcasts.— The  Lehigh  Valley  Coal  Co.  in  1907,  started  sub- 
stituting for  the  wood  overcast  one  of  concrete  and  steel.  These 
overcasts  will  be  permanent  and  substantial,  their  destruction 
only  being  accomplished  by  a  squeeze,  in  which  event  all  other 
construction  would  be  destroyed  as  well. 

The  construction  of  the  overcasts  is  shown  in  Fig.  11.  Being 
located  in  an  old  portion  of  the  mine,  much  gob  and  other 


484 


COAL  MINING  COSTS 


g  i  _     _L 

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FIG.  11.— Plan  View. 


MISCELLANEOUS  INSIDE  COSTS 


485 


H    S 


486  COAL  MINING  COSTS 

refuse  is  found  011  both  sides  of  the  road  and  must  be  cleared 
away  at  the  immediate  location  of  the  bridges.  After  this 
trenches  for  the  masonry  walls  to  carry  the  bridge  are  dug 
and  carried  to  sufficient  depth  below  the  bottom  slate  of  the 
vein  to  obtain  a  solid  foundation.  The  walls  are  then  built 
4  ft.  thick,  consisting  of  rock  and  bone  of  sufficient  size  selected 
by  the  mason  from  the  gobbed  refuse  made  and  packed  in  the 
chambers  during  mining.  Lime  mortar  is  used  and  the  faces 
of  the  wall  are  almost  entirely  surfaced  off  with  a  coat  of 
mortar  after  completion. 

Next  in  order,  timbers  for  carrying  the  concrete  are  placed 
by  the  timberman,  and  consist  of  second-hand  3  X  5-in.  wooden 
rails  taken  from  old  chambers  and  of  second-hand  mine  ties 
and  props  for  uprights,  on  which  are  laid  2-in.  planks  double 
thick,  which  in  turn  carry  2  in.  planks  on  edge.  To  these  planks 
are  nailed  the  2-in.  planks  which  carry  the  concrete. 

The  time  consumed  is  somewhat  greater  than  one  might  con- 
sider and  an  explanation  is  necessary.  The  clearing  away  of 
the  refuse  accumulated  is  very  laborious  and  progress  is  slow, 
and  in  excavating  for  the  trenches  considerable  heavy  rock 
and  bone  must  be  removed.  The  work  being  done  in  the  return 
current  makes  it  warm  for  men  to  work,  and  less  is  accom- 
plished than  if  the  same  were  located  in  a  fresh-air  current. 
Again,  the  rock  and  bone  constituting  the  material  for  the 
masonry  wall  are  all  collected  in  the  old  chambers  in  the 
vicinity  and  taken  to  the  location  for  the  wall.  The  broken 
stone  used  in  the  concreting  is  obtained  from  a  chamber  some 
distance  away,  the  stone  having  been  dumped  there  during 
the  driving  of  the  rock  slope.  The  mason  selects  what  material 
is  fit  for  use.  At  the  same  time  sand  suitable  for  concreting 
is  gathered  from  the  same  place,  having  been  made  during  the 
blasting  in  the  slope  and  loaded  out  with  the  rock. 

As  will  be  seen  by  the  end  view,  Fig.  11,  two  concrete 
walls  form  the  sides  of  the  overcast  and  become  necessary 
where  the  entire  vein  has  been  mined  out.  The  Baltimore 
vein,  which  in  this  particular  location  is  one  seam,  more  fre- 
quently splits  into  two  distinct  seams,  the  top  split  known  as 
the  Cooper  or  Upper  Baltimore,  and  the  bottom  split  known 
as  the  Bennett  or  Lower  Baltimore  vein.  At  the  location  in 
question  the  dividing  slate  is  only  about  1  ft.  thick  and  in 


MISCELLANEOUS  INSIDE  COSTS  487 

most  instances  both  splits  have  been  removed.  At  the  location 
of  No.  4  overcast  only  the  bottom  split  was  mined  and  it  was 
necessary  to  take  down  the  top  vein  to  get  sufficient  height 
for  the  roof  of  the  overcast.  In  taking  down  this  top  coal,  or 
split,  the  ribs  were  neatly  dressed  and  the  concrete  floor  of 
the  overcast  was  carried  high  enough  to  dispense  with  the 
walls.  At  the  location  for  Nos.  1,  2  and  3  overcasts  the  total 
vein  has  been  mined  and  concrete  side  walls  will  become  neces- 
sary. 

The  two  masonry  walls  are  built  4  ft.  thick,  the  walls  on  the 
up-pitch  side  being  on  an  average  7  ft.  high,  the  one  on  the 
down-pitch  side  averaging  10  ft.  high. 

The  concrete  floor  is  12  in.  in  thickness  and  22  ft.  long  b> 
20  ft.  wide,  consisting  of  a  1-2-3  mixture.  The  breadth  and 
length  will  of  course  be  greater  or  less  depending  on  the 
dimensions  of  the  openings  through  the  coal  at  the  different 
locations. 

The  T-iron  rails  are  spaced  12  in.  center  to  center,  the  bot- 
tom or  base  of  the  rail  being  embedded  2  in.  in  the  concrete. 
During  concreting  the  rails  are  supported  by  blocking  under 
the  head. 

For  the  side  walls  a  1-2-5  mixture  will  be  used,  the  walls 
being  12  in.  thick  and  cement  worked  well  into  the  crevices  of 
the  coal  and  top  rock. 

The  details  of  cost  of  the  No.  4  overcast  are  given  here- 
with, the  figures  being  as  of  1907.  The  day's  work  consists  of 
nine  hours  and  the  hourly  rate  includes  the  strike  percentages 
and  sliding  scale.  The  proportions  for  the  concrete  mixture 
were  1-2-3.  Sand  and  stone  for  the  concrete  and  walls  do  not 
appear  in  cost  for  material,  the  only  cost  on  these  items  being 
included  in  the  labor. 

The  total  cost  per  cubic  yard  for  the  masonry  wall  is  there- 
fore $1.63.  The  cost  for  labor  and  material  is  about  equal  for 
the  overcast  or  $5.25  per  cu.yd.  for  each,  a  total  of  $10.50  per 
cu.yd.  The  two  walls  needed  for  Nos.  1,  2  and  3  overcasts 
make  an  additional  11.5  cu.yd.  at  a  somewhat  smaller  rate 
per  cu.yd.,  say  $8,  so  that  the  entire  cost  of  the  overcast, 
including  two  supporting  masonry  walls  and  two  concrete  side 
walls,  in  a  vein  of  this  thickness  will  be  about  $365,  and  about 
$270  without  the  side  walls. 


488  COAL  MINING  COSTS 

MASONRY  WALLS,  60.5  Cu.  YD. 
LABOR 

Clearing  away  refuse,  digging  trenches,  getting  stone, 
mixing  mortar  and  building  two  walls,  2  masons,  20 
days  each,  or  360  hr.  at  25.4c $91.44 

MATERIAL 

25  bu.  of  lime  at  30c . .  7 . 50 


$98.94 
CONCRETE  OVERCAST,  16.5  CU.YD. 

LABOR 

Getting  lumber  for  supporting  concrete  work  and  placing 

same,  3  men  at  2  days  each,  or  54  hr.  at  25.2c $13.61 

Placing  2-in.  plank  for  concreting,  1  man  2  days  or  18  hr. 

at  25.2c 4.54 

Getting  broken  stone,  mixing  and  placing  concrete,  2 

masons  at  14  days  each  or  252  hr.  at  25 .2c 63.50 

Bending  and  transporting  rails  approximate 5 . 00 


$86.65 

MATERIAL 

120  bags  portland  cement  at  45c $54 . 00 

1000  ft.  2-in.  hemlock  plank  at  22c 22 . 00 

20  Ib.  20d.  nails  at  2£c 50 

1850  Ib.  (T%-  ton)  second-hand  25-lb.  T-iron  rails  at  $12.50 

a  ton .  .  .   10 . 00 


$86.50 

Tile  stoppings  and  overcasts. — Terra-eotta  blocks  have 
proven  their  usefulness  in  a  mine,  as  well  as  for  the  purposes 
to  which  they  are  put  outside.  Not  only  have  they  been  used 
economically  for  stoppings  along  the  main  air  courses,  but  they 
are  now  being  adopted  successfully  for  the  walls  of  overcasts. 
It  appears  also  that  scrap  iron  is  to  be  displaced  by  a  rein- 
forcement which  gives  a  concrete  structure  not  only  more 
artistic  but  more  economical  and  more  quickly  erected. 

The  Consolidation  Coal  Co.,  as  a  preventative  of  fire  at  the 
same  time  as  a  safety-first  measure,  decided  about  1914  to 
eliminate  wood  stoppings  and  wood  overcasts  where  practi- 
cable and  to  allow  no  more  new  overcasts  anywhere  to  be  con- 
structed of  combustible  material,  the  same  to  apply  to  perma- 
nent stoppings  along  main  entries. 


MISCELLANEOUS  INSIDE  COSTS  489 

The  most  natural  step  was  to  use  the  cinders  from  the 
boiler  house  at  the  mine  and  to  build  these  structures  of  con- 
crete. However,  investigation  proved  that  too  much  money 
was  being  spent  for  these  improvements.  The  stoppings  were 
reduced  to  a  thickness  of  4  in.,  but  the  cost  still  appeared  too 
large. 

The  terra-cotta  blocks  were  first  used  for  stoppings  and 
their  success  was  readily  apparent.  As  the  4-in.  cinder-concrete 
stopping  had  proven  stable,  a  hollow  block  of  4-in.  width  was 
at  first  advocated ;  but  the  5  X  8  X  12-in-  block  was  finally 
adopted. 

On  main  entries  where  the  stoppings  are  intended  to  remain 
for  a  considerable  length  of  time  they  are  bonded  with  a  mortar 
composed  of  one  part  cement  to  two  parts  sand.  On  room 
headings  where  these  stoppings  have  been  built,  lime  was  used 
in  place  of  the  cement,  this  making  it  possible  to  tear  down  the 
stopping  when  desired,  without  destroying  the  tile. 

Sufficient  cost  data  have  been  obtained  on  the  construction 
of  this  type  of  stopping  at  the  mines  using  this  material  to 
show  a  saving  of  from  40  to  50  per  cent  in  every  instance  as 
against  the  use  of  concrete.  The  cost  compared  to  wood  stop- 
pings for  room  headings  shows  about  the  same  difference  in 
favor  of  the  wood.  It  was  thought  that  by  the  recovery  of  the 
blocks  and  their  reuse  this  difference  could  be  more  than  off- 
set; but  accurate  costs  kept  on  this  work  shows  that  it  would 
be  necessary  to  recover  these. blocks  several  times  over  to  do 
this.  Therefore,  it  is  not  advisable  to  use  the  blocks  for  tem- 
porary structures  unless  for  other  reasons  than  economy. 

The  crosscuts  in  the  mines  of  the  Fairmont  region  run  about 
10  to  12  ft.  wide  and  7  to  8  ft.  high,  and  average  about  80 
sq.ft.  in  area.  It  has  been  found  that  two  men  will  construct 
three  or  four  stoppings  per  day,  providing  the  material  is 
placed  ready  for  their  use. 

It  might  be  possible  to  reduce  the  cost  of  stoppings  further 
by  the  use  of  ' ' Self-sentering, "  "Hy-Rib,"  or  "Rib-Lath," 
plastering  these  with  a  cement  mortar  an  inch  or  two  thick. 
A  stopping  made  with  any  of  these  materials  properly  placed, 
should  carry  as  much  as,  or  more  pressure  than,  the  ordinary 
wood  stopping.  The  plan  of  erection  would  be  simple,  con- 
sisting of  trenching  the  ribs,  roof  and  bottom  an  inch  or  two, 


490  COAL  MINING  COSTS 

slipping  the  adopted  metal  in  place  and  plastering  with  cement 
or  lime  mortar,  the  material  depending  upon  the  desired  life 
of  the  stopping.  Price  and  the  strength  which  could  be  secured 
should  govern  the  selection  of  the  reinforcement.  A  heaving 
of  the  bottom  would  in  all  probability  be  cut  by  the  stopping, 
since  heaving  is  caused  by  the  swelling  of  the  fireclay,  this 
material  at  the  same  time  softening  markedly.  If  the  pressure 
comes  from  the  top  or  roof  there  might  be  much  deflection 
in  the  stopping  without  doing  it  any  material  injury,  besides 
it  could  be  as  easily  repaired  as  the  wood  stopping  in  case  of 
sufficient  pressure  to  cause  cracks  in  the  plaster. 

All  well  ventilated  mines  must  contain  overcasts,  and  the 
costs  of  these  structures  soon  run  into  considerable  amounts, 
if  they  are  not  carefully  planned  and  constructed.  There  are 
few  mine  foremen  or  even  superintendents  who  take  the  trouble 
to  ascertain  accurately  just  what  it  costs  to  erect  an  overcast 
so  that  he  can  tell  you  how  much  was  spent  per  cubic  yard 
for  concrete  or  what  the  walls  and  roof  cost  separately. 

Fig.  12  shows  the  type  of  overcast  now  being  constructed 
in  the  mines  of  The  Consolidation  Coal  Co.  in  the  Fairmont 
Region.  Blank  spaces  are  left  in  the  blueprints,  and  dimensions 
to  suit  the  location  are  supplied.  Since  the  sheets  of  "Self- 
sentering"  are  sold  in  even  foot  lengths,  the  distances  between 
walls  are  made  to  correspond.  Thus  for  an  11-ft.  sheet  2  in. 
are  deducted  for  the  12-in.  rise  and  two  or  more  inches  on 
each  side  for  bearing  spaces,  leaving  the  distance  between 
walls  10  ft.  6  in. 

This  style  of  overcast  makes  use  of  the  terra-cotta  blocks, 
the  5  X  8  X  12-in.  block  being  laid  to  give  8-in.  walls.  On 
top  of  each  wall  is  placed  a  layer  of  concrete  on  which  is  set 
a  small  steel  rail  2  in.  from  the  edge  of  the  wall.  Then  the 
reinforcing  sheets  of  ' '  Self-sentering, "  which  come  already 
shaped,  are  placed  abutting  on  the  flanges  of  the  rails  and  the 
center  line  or  middle  of  the  reinforcement  supported  by  a 
3  X  5-in.  wood  stringer ;  then  the  concrete  is  spread  on  the  top 
as  required  on  the  plan.  After  the  concrete  sets,  the  wood 
stringer  is  knocked  down  and  a  coat  of  cement  mortar  plas- 
tered underneath.  Not  a  single  wood  form  is  required  on  the 
whole  structure,  except  around  the  iron  rails  on  top  of  the 
w^alls,  and  even  this  can  be  eliminated  by  standing  a  row  of 


MISCELLANEOUS  INSIDE  COSTS 


491 


the  terra-cotta  blocks  on  the  5-in.  side  just  outside  of  the  rails, 
thus  making  them  serve  as  forms. 


f  Cinder  concrete.  /••?••/•  cinders  to  be 
thoroughly  netted,  concrete  mined 
fo  the  consistency  of  paste  and  placed 
on  "&elf-sente  ring  "remforcement 

&X..5" posts  nitna3xS"cap 
•fo  be  placed  on  center  line 
to  support  arch  while 
concrete  roof  is  being 
placed 

\. 

[Self-sentering* 
reinforcemenr 
Ho.  24  gage 


(  3  Concrete  on  top  of  rt in  forcf- 
<  ment,  I  coat-cement  mortar 
(.  underneath 


FIG.  12. — Reinforced  concrete  overcast  supported  on  terra-cotta  blocks  used 
by  the  Consolidation  Coal  Co. 

There  are  times  when  it  is  necessary  to  deviate  from  the 
arched-roof  type  on  account  of  the  skew  of  the  air  bridge,  in 
which  case  it  is  more  economical  to  use  the  flat  roof.  Fig.  13 


x5  Stringer 
entire  length 
ofovercasr 


I" Cement  footing      each  pipe 
SECTION  A.-A 


FIG.    13. — The    Consolidation    Coal   Co.'s   reinforced   concrete,    hollow   tile 
overcast  with  flat  roof. 

presents  a  design  of  this  kind  where  the  roof  is  supported  by 
three  or  four  old  boiler  tubes,  the  reinforcement  being  sus- 
pended from  these  tubes  or  pipes  by  means  of  wire  of  sufficient 


492  COAL  MINING  COSTS 

strength  to  carry  the  weight.  This  type  of  overcast  also  needs 
only  temporary  supports  along  the  middle  while  the  concrete 
is  being  placed,  thus  saving  considerable  labor  and  material  in 
the  construction  and  emplacement  of  forms. 

The  Utah  Fuel  Co.  constructed  six  concrete  stoppings  and 
one  overcast  in  some  work  they  did  about  1914.  On  account 
of  the  height  of  the  stoppings  they  were  built  of  reinforced 
concrete  with  reinforcing  wings  as  shown  in  the  accompanying 
drawings,  Figs.  14  and  15,  old  rails  and  wire  rope  being  used 
for  the  reinforcing  material.  The  reinforcement  of  the  over- 
cast was  made  with  the  same  class  of  material.  The  cost  of 
these  is  given  in  the  figures. 

Comparison  of  doors  and  overcasts. — A  mine  door  is  a 
costly  as  well  as  dangerous  item  of  equipment,  yet  this  seldom 
receives  the  thought  that  it  deserves.  When  it  is  necessary  to 
install  a  permanent  door  the  conditions  should  be  carefully 
considered  to  see  if  it  will  not  be  cheaper  and  safer  to  put  in 
an  overcast. 

A  permanent  door  involves  much  expenditure  in  addition  to 
that  for  lumber.  The  wages  of  two  men  must  be  paid  for  its 
construction  and  erection.  A  trapper  will  have  to  be  con- 
stantly employed  to  open  and  close  it.  A  shelter  hole  must 
be  constructed  at  the  door  and  another  hole  also  must  be 
provided  for  a  barrel  of  water,  for  all  permanent  doors  should 
be  protected  against  a  possible  fire.  The  continual  breaking 
and  smashing  of  the  door  by  trips  involves  a  further  expendi- 
ture for  lumber  and  for  the  workmen  who  replace  it. 

A  light  should  be  provided  at  all  doors.  This  involves  the 
wiring  of  the  light,  the  replacing  of  globes,  and  the  employ- 
ment of  wire  men  who  must  spend  a  day,  more  or  less,  in 
installing  it.  Trolley  wires  must  be  guarded  on  both  sides  of 
the  door,  which  means  the  use  of  more  lumber  and  the  employ- 
ment of  more  labor.  These  boards  often  are  knocked  down 
by  trolley  poles  and  again  the  service  of  a  high-class  or  highly- 
paid  man  is  needed. 

Where  permanent  doors  are  used  an  extra  door  should  be 
placed  that  will  have  the  same  effect  on  the  ventilation.  This 
door,  of  course,  should  be  a  full  trip  length  away.  Heavy 
blasting  or  heavy  caving  often  damages  doors.  The  weighting 


MISCELLANEOUS  INSIDE  COSTS 


493 


of  the  roof,  the  movement  of  the  sides  or  the  heaving  of  the 
bottom  will  affect  them  also. 

Looking  at  the  matter  from  a  safety  viewpoint  we  often 
find  the  door  frame  cuts  the  clearance,  frequently  to  as  little 


30-lb.  T  Rail  Reinforcing, 
'     J5  C+oC. 


ELEVATION 

FIG.  14. — Plan  and  elevation  of  the  Utah  Fuel  Co.'s  concrete  overcast. 


.--.26-0*- - >i  2  ft-       |< M-O- 

OOUBLE  BRACED  STOPPING 

DETAIL  OF  COST  PER  CU.  YARD 

Stoppings  and  I  overcast  Total  119.0  Cubic  Yards 
d  work  Labor  $3.15  per  8Hrs.  Carpenters$3.40  perSHr* 


ITEM 

LABOR 

1ATERIAU 

TOTAL 

Reinforcinq  Material  (Scrap  Iron) 

i>  0.278 

$0.278 

Qatherina  &  Distributing  Gravel  in  Mine 

2.683 

Cement 

0.076 

$2.659 

2  735 

Forms 

1.936 

0.554- 

2490 

Redistribute  Gravel.Mix  &  place  Concrete 

3.835  * 

3.835 

Totals. 

&  e.&oa 

&3.ZI3 

fT2.02T— 

Gravel  hauled  in  winter. Difficult  to  redistribute  gravel  and  place 
concrete  on  account  of  old  workings 

*  INCLUDES  COST  OF  MAKING  OLD  WORKINGS  SAFE 

FIG.  15. — Plan  and  cost  of  reinforced-concrete  stoppings  at  the  Utah  Fuel  Co.'s 

mines. 

as  12  in.  A  brakeman  or  driver,  knowing  he  has  30  in.  clear- 
ance, forgets  the  door  frame  is  in  his  way  and  accordingly 
at  this  point  a  man  may  be  squeezed  to  death.  The  required 
shelter  hole  is  not  always  provided,  and  an  instance  is  recorded 


494  COAL  MINING  COSTS 

where  a  trapper  stood  back  of  his  door  and  the  heavy  iron 
guard  receiving  the  weight  of  the  cars  crushed  him  so  severely 
that  he  died. 

Should  the  boy  neglect  his  work  the  motorman  is  in  great 
risk  of  losing  his  life  by  running  into  the  closed  door.  The 
cross  bar  over  the  door  often  is  far  below  the  uniform  height 
of  the  roof  and  may  easily  dash  a  man's  brains  out.  Doors 
have  been  known  to  take  fire  and  cause  serious  damage.  One 
at  Delagua,  Col.,  is  thought  to  have  been  set  on  fire  by  the 
trapper  and  been  the  cause  of  a  destructive  fire  in  which  many 
lives  were  lost. 

Trappers,  drivers  and  workmen  leave  doors  fastened  or 
thrown  back  and  as  a  result  that  portion  of  the  mine  where 
the  air  is  short-circuited  may  be  endangered  by  an  accumula- 
tion of  gas,  with  a  result  that  is  not  to  be  reckoned  in  dollars 
and  cents  but  in  humanity.  It  is  not  advisable  to  place  doors 
at  the  foot  of  steep  grades.  Even  automatic  doors  are  danger- 
ous where  a  man  cannot  control  his  trip,  as  these  doors  are 
not  always  positive  in  action. 

The  penalty  under  the  compensation  schedule  for  neglect 
to  dispense  with  the  door  certainly  makes  it  a  costly  and 
dubious  economy.  Doors  must  be  so  hung  that  they  will  close 
themselves,  must  be  strongly  built,  tightly  sealed  into  the  roof, 
and  a  second  door  that  has  the  same  effect  on  the  ventilation 
is  required. 

For  neglecting  to  comply  with  these  specifications  a  charge 
of  Ic.  per  $100  of  payroll  is  the  penalty  provided.  Under  the 
rules  of  the  compensation  schedule  shelter  holes  must  be  main- 
tained at  all  permanent  doors.  A  penalty  of  6c.  can  be  charged 
the  operator  not  complying  with  this  requirement.  Doors  must 
be  whitened  or  enclosed  lights  maintained.  A  6c.  penalty  is 
provided  for  non-compliance  with  the  requirement  that  the 
trolley  wire  at  doors  be  guarded  on  both  sides. 

A  clearance  of  30  in.  must  be  provided  between  the  widest 
part  of  the  motor  or  cars  and  the  frame  of  the  door  or  a  12c 
charge  can  be  made.  When  these  charges  are  totaled  we  find 
that  one  door  can  create  a  maximum  of  31c.  per  $100  payroll. 
When  the  total  payroll  of  the  mine  is  computed  and  this  pro- 
portion deducted  it  will  give  some  idea  why  a  permanent  door 
should  not  be  placed,  but  even  when  conditions  seem  to  warrant 


MISCELLANEOUS  INSIDE  COSTS  495 

paying  the  penalty  to  save  the  first  cost  of  an  overcast  it  is 
generally  more  advisable  to  erect  it. 

The  initial  cost  is  greater,  of  course,  but  in  the  end  the 
overcast  will  be  found  to  have  paid  for  itself  in  saving  the 
expenditures  enumerated.  With  the  overcast  the  clearance 
can  be  made  ample. 

Cost  of  stone  and  wood  brattices. — An  ordinary  wooden 
brattice,  single  thickness,  may  be  figured  as  follows,  figures  as 
of  1907 : 

1-in.  plank,  £-m.  strips  and  waste,  100ft.  B.M.  @  $20  per 

thousand $2 . 00 

Labor  $1.75,  3  props  and  caps,  5c 1 .90 

Daubing  35c.,  nails,  etc.,  25c 0 . 60 


$4.50 

The  life  of  this  brattice  will  ordinarily  be  from  4  to  5  yr., 
so  that  $1  per  year  per  brattice  may  be  taken  as  the  cost  of 
maintenance.  Where  top  is  shot,  or  where  there  is  much  draw 
slate,  it  will  usually  be  found,  particularly  in  low  coal,  that 
there  is  considerable  cleaning  out  to  do  to  get  at  the  old  brat- 
tices; which  work  may  readily  cost  as  much  as  the  brattice 
itself  thus  making  the  cost  of  maintenance  $2  per  year,  instead 
of  $1. 

Doors  cost  in  1907: 

Lumber,  say  120  ft.  B.M.  @  $20,  $2.40,  nails  $0.10 $2.50 

Labor  $3.75,  hinges  $0.75 4.50 


$7.00 

There  are  in  most  mines,  at  least  one  or  two  places  where 
a  door  boy  at  75c.  per  day,  or  say  $200  per  year,  could  be 
replaced  by  an  overcast  at  $50  first  cost,  good  for  5  yr.,  and 
thus  costing  only  $10  per  year,  making  a  net  saving  of  $190 
per  year  for  each  door  boy  replaced. 

Where  the  old  brattices  have  to  be  dug  out  before  they 
can  be  replaced,  this  work  may  readily  amount  to  $1  per  year 
for  each  old  brattice,  or  $90  additional.  Where  the  mine  work- 
ings are  so  arranged  that  the  brattices  on  the  cross  entries  will 
not  be  in  service  longer  than  the  life  of  the  first  brattices  put 
in,  the  $160  for  maintenance  may  be  dropped.  It  often  occurs 
that  shorter  new  lines  are  driven,  but  this  is  objectionable,  as 


496 


COAL  MINING  COSTS 


it  leaves  no  suitable  air  current*  along  the  haul  ways  to  carry 
off  the  dust  from  haulage  and  this  is  the  class  of  dust  most  to 
be  dreaded. 

Where  the  brattices  have  to  be  dug  out  the  total  cost  may 
be  $2250,  or  1.25c.  per  ton,  for  a  year 's  output. 

Stone  brattices  are  preferably  from  1^  to  2y2  ft.  thick, 
depending  upon  thickness  of  coal,  top,  etc.,  and  should  be  well 
mined  or  cut  into  the  rib,  particularly  in  the  case  of  soft  coals 
that  spall  off  readily.  Where  the  coal  is  low  and  hard,  with  good 
roof  and  bottom,  a  1-ft.  wall  would  seem  ample.  The  follow- 
ing table  shows  the  contents  in  cubic  yards  of  different  sizes 
of  brattices: 


FOUR-FOOT  COAL 


Width  Crosscuts 
or  Breakthroughs  in 

1  Foot  Thick, 

1.5  Foot  Thick, 

2  Feet  Thick, 

Feet 

Cubic  Yards 

Cubic  Yards 

Cubic  Yards 

10.0 

1.48 

2.22 

2.96 

12.0 

1.78 

2.67 

3.56 

14.0 

2.08 

3.11 

4.15 

16.0 

2.37 

3.56 

4.74 

18.0 

2.67 

4.00 

5.33 

20.0 

2.96 

4.44 

5.92 

22.0 

4.89 

6.52 

8.15 

24.0 

3.56 

5.33 

7.11 

SIX-FOOT  COAL 


Width  Crosscuts  in 

1  Foot  Thick, 

1.5  Feet  Thick, 

2  Feet  Thick, 

Feet 

Cubic  Yards 

Cubic  Yards 

Cubic  Yards 

10.0 

2.22 

3.33 

4.44 

12.0 

2.67 

4.00 

5.33 

14.0 

3.11 

4.67 

6.22 

16.0 

3.56 

5.33 

7.11 

18.0 

4.00 

6.00 

8.00 

20.0 

4.44 

6.67 

8.89 

22.0 

4.89 

7.33 

9.78 

24.0 

5.33 

8.00 

10.67 

MISCELLANEOUS  INSIDE  COSTS  497 

Where  the  draw  slate  forms  a  durable  building  stone, 
double-stone  brattices,  filled  with  muck,  may  be  built  for  $10 
to  $15  (figures  as  of  1907),  part  of  the  actual  cost  being  prop- 
erly chargeable  to  " slate,"  or  entry  cleaning.  Where  stone 
must  be  quarried  and  brought  in  from  the  outside,  a  good 
brattice,  single  thickness,  may  cost  anywhere  from  $15  to  $35. 
Where  coke  cinder  is  available,  cinder  concrete  may  be  used 
for  brattices  and  overcasts. 

Disregarding  the  cross  entries,  which  may  be  assumed  to 
be  the  same  in  either  case,  and  taking  $25  as  the  cost  of  a 
stone  brattice,  which,  particularly  for  low  coal,  may  be  con- 
sidered a  liberal  estimate,  the  costs  of  wooden  and  stone  brat- 
tices may  be  figured  as  follows: 

Stone  brattices: 

Main,  10  brattices  @  $25.00  per  year $250.00 

Wood  brattices: 

Main,     90     old     brattices,     maintenance 

@  $1.00 $90.00 

Main,  10  new  brattices  @  $4.50 45 . 00 

135.00 


$115.00 

The  balance  of  $115  in  favor  of  wood  brattices  is  equivalent 
to  0.064c.  per  ton,  for  180,000  tons. 

Figuring  the  cost  of  maintenance  of  wooden  brattices  at 
$2,  instead  of  $1,  the  balance  in  favor  of  wood  brattices  is  only 
$25,  or  equivalent  to  0.014c.  per  ton  for  180,000  tons. 

Figuring  stone  brattices  at  $25,  and  the  maintenance  of 
wooden  at  $1,  a  stone  brattice  will  need  to  be  in  service  25  yr. 
to  be  as  cheap  as  a  wood  brattice ;  but  figuring  the  maintenance 
at  $2,  will  only  need  to  be  in  service  12^  yr. 

Double  wood  brattices  filled  with  muck  may  be  figured  at 
$10  to  $11  each,  with  say  $2  to  $2.25  per  year,  for  maintenance. 
Compared  with  this  a  stone  brattice  will  only  need  to  be  in 
service  12%  yr.,  or  allowing  $1  per  year  for  digging,  say  8  yr.. 
to  be  as  cheap  as  a  wooden  brattice.  Stone  brattices  cost 
somewhat  more  than  the  double  air-course  arrangement,  where 
the  latter  does  not  involve  shooting  considerable  top,  or  equiva- 
lent yardage  for  narrow  work,  but  cost  less  where  yardage 
must  be  paid. 


498  COAL  MINING  COSTS 

Refuge  Chambers. — Subsequent  to  the  Cherry  (Illinois) 
mine  fire  there  was  a  general  feeling  among  the  engineers  of 
the  country  that  underground  refuge  chambers  should  be 
established  at  all  mines  to  prevent  a  repetition  of  this  insofar 
as  was  humanely  possible.  A  paper  was  presented  before  the 
West  Virginia  Mining  Institute  in  1910,  dealing  with  this  ques- 
tion some  excerpts  from  which  are  given  herewith. 

The  maximum  size  of  a  district  to  be  supplied  by  a  refuge 
chamber  depends  somewhat  on  the  geological  and  other  phys- 
ical conditions  presented  by  the  seam  and  the  system  of  work- 
ing same.  It  would  seem  desirable  to  have  it  bear  some  rela- 
tion to  the  maximum  number  of  men  employed  in  a  district 
ventilated  by  a  separate  split  of  air.  We  will  assume  that  the 
maximum  number  of  men  is  one  hundred,  a  not  uncommon 
maximum  allowed  for  a  single  split  of  air.  As  there  will  be 
new  districts  or  panels  forming  while  others  are  being  worked 
out,  the  average  number  of  men  we  will  figure  at  50.  A 
medium-sized  mine  has  about  200  men  employed  on  the  day 
shift,  and  a  large  mine  about  500.  Accepting  the  average  of 
50  men  in  a  district,  there  would  be  from  4  to  10  live  districts 
in  a  medium-  to  large-sized  mine,  and  as  many  refuge  chambers 
under  the  system  proposed. 

COST  OF  REFUGE  CHAMBER 

500  ft.  2-in.  common  pipe  casing,  in  place,  say $50 

50  ft.  of  excess  room  neck  yardage  and  special  entrance,  say    50 
5  room  crosscuts,  say,  100  ft.  of  yardage 50 

5  masonry  stoppings,  at  $10 50 

6  masonry  door  frames,  at  $5 30 

6  doors  and  frames,  at  $6 36 

Sanitary  closet  and  fixtures 15 

Wall  cases  with  glass  fronts 20 

Casks,  pails  and  miscellaneous  fittings 10 

Food  in  tins  and  cans,  say 25 

6  dry  cell  electric  lights,  say  $5  each 30 

2  safety  lamps,  at  $5 10 

1  oxygen  resuscitating  box,  with  two  cylinders 45 

First  aid  box,  medicines  and  disinfectants 25 

Miscellaneous,  say 54 

Total $500 

To  establish  these  refuge  chambers  may  appear  to  be  a 
serious  task,  but  if  they  are  planned  for  in  laying  out  the  mine, 


MISCELLANEOUS  INSIDE  COSTS  499 

the  cost  per  ton  would  be  insignificant.  Nearly  all  modern 
coal  developments,  as  a  matter  of  good  engineering,  are,  or 
should  be  preceded  by  thorough  prospecting,  both  to  knoiv  the 
continuity  of  the  seams  and  to  properly  plan  the  mine. 

In  the  following  estimate,  the  room  is  not  considered  an 
added  expense,  except  for  the  extra  length  of  room  neck.  The 
cost  of  drilling  the  hole  is  considered  part  of  the  cost  of  pros- 
pecting; the  cost  of  its  casing  for  an  assumed  depth  of  500  ft. 
is  alone  considered.  The  telephone  is  not  regarded  as  an  extra 
cost. 

The  foregoing  provides  for  a  good  equipment ;  other  appara- 
tus mentioned  previously  should  be  considered  as  part  of  the 
mine  equipment. 

If  a  mine  had  6  such  stations,  the  cost  underground  would 
be  $3000.  On  the  surface  the  special  equipment  would  vary 
widely  with  the  physical  conditions  and  regular  equipment. 
If  a  mine  used  compressed  air,  the  only  additional  cost  for  the 
stations  would  be  the  outside  pipe  lines.  These  pipe  lines  need 
not  be  large,  as  economy  of  operation  would  not  enter  into  the 
calculations.  It  is  probable  that  all  such  lines  to  drill  holes  of 
six  refuge  chambers  could  be  supplied  at  from  $1500  to  $2000, 
under  ordinary  conditions. 

When  the  mine  has  an  electric  plant  but  not  a  compressor 
plant,  the  additional  surface  equipment  would  be  the  cost  of 
the  power  lines  to  the  various  drill  holes  and  the  cost  of  the 
small  motor-driven  fans  or  compressors.  Each  drill  hole  sur- 
face instalment  could  probably  be  put  in  at  a  cost  not  exceeding 
$500. 

When  a  mine  had  neither  compressed-air  nor  electric  plant, 
the  cost  of  instalment  would,  of  course,  be  much  greater,  as  it 
would  involve  a  small  central  plant.  However,  it  may  be 
pointed  out  that  such  a  plant  would  be  extremely  useful,  and 
no  doubt  pay  for  instalment  on  other  grounds. 

Let  us  assume  that  the  average  total  cost  of  instalment  of 
district  refuge  chambers  figures  as  much  as  $10,000,  or  let 
us  say  5  per  cent  of  the  total  cost  of  the  mine  investment,  the 
possibility  of  saving  a  considerable  number  of  lives,  if  disaster 
comes,  makes  it  seem  a  good  investment. 

Mine  sprinkling  costs.— At  the  Sunnyside,  No.  2  Mine  in 
Carbon  County,  Utah,  quite  an  elaborate  sprinkling  system 


500  COAL  MINING  COSTS 

was  installed  about  1908.     In  this  system  the  smallest  pipe 
used  on  the  haulageways  was  1%  in.  and  in  the  rooms,  %  or 

1  in.  pipe  is  used  equipped  with  a  brass  hose  bibb  and  kept 
within  200  ft.  of  the  working  face. 

In  operation,  two  men  are  continually  employed  and  they 
are  required  to  attend  to  all  extensions  and  repairs,  as  well  as 
keep  the  mine  sprinkled.  In  an  ordinary  mine  the  necessary 
work  on  pipe  lines  will  not  occupy  more  than  an  average  of 

2  or  3  hr.  of  their  time  per  day,  the  work  including  extensions 
in  rooms  and  entries  where  work  is  advancing,  taking  up  pipe 
where  work  is  retreating  and  roof  expected  to  cave,  repair- 
ing broken  pipe,  packing  leaky  valves,  etc. 

Each  water  man  will  carry  with  him  150  to  200  ft.  of  %-in. 
rubber  hose,  attaching  one  end  to  the  hose  bibbs  which  are 
opened  by  the  key  or  lever,  which  he  carries.  He  will  thor- 
oughly wet  down  the  roof,  floor,  and  sides,  or  ribs,  of  all  open- 
ings accessible,  paying  especial  attention  to  the  vicinity  of 
working  faces,  brattice  and  timbers  and  from  time  to  time 
wetting  down  abandoned  rooms,  etc.,  which  are  still  open. 

The  water  men  are  instructed  to  keep  all  parts  of  the  mine 
sufficiently  damp  so  that  upon  taking  a  handful  of  debris  from 
the  floor,  and  subjecting  it  to  pressure  of  the  hand,  it  will 
cake  and  retain  its  shape  after  removal  of  pressure.  All  brat- 
tice must  also  be  kept  damp,  this  requirement  alone  demand- 
ing the  presence  of  water  men  at  least  every  other  day.  If 
fine  dust  is  dampened  and  then  allowed  to  dry,  it  is  very  diffi- 
cult to  penetrate  it  with  water,  due  to  the  tendency  of  this 
fine  dust  to  form  an  almost  impervious  film  of  dust  around 
globules  of  water,  while  if  this  dust  is  kept  damp,  there  is  no 
dry  fine  dust  present  to  imprison  and  waste  the  water.  Pour- 
ing water  in  quantity,  as  from  pail  or  barrel,  on  dry  fine  dust, 
is  wasteful  of  water  and  absolutely  ineffective,  due  to  the 
tendency  of  the  fine  dust  to  form  the  film  above  mentioned, 
while  the  use  at  frequent  stated  intervals  of  a  stream  with 
good  pressure  finely  divided  by  hose  or  nozzle  moves  the  dust 
and  permits  the  water  to  penetrate  the  dust  particles,  not 
superficially,  but  through  to  the  floor. 

At  Sunny  side  No.  2  Mine,  Carbon  County,  Utah,  the  coal 
vein  is  6  to  10  ft.  thick,  and  pitches  about  10  per  cent.  Here 
the  coal  is  absolutely  dry  and  dust  very  inflammable,  making 


MISCELLANEOUS  INSIDE  COSTS  501 

the  wetting  of  dust  absolutely  necessary.  The  above-described 
sprinkling  system  is  here  an  unqualified  success.  The  cost  of 
the  system  was  about  as  follows  in  1907 : 


Labor 

Materials 

Total 

50,000  gal.  redwood  tank  in  place  
Pressure  pump  in  powerhouse 

$600 
175 

$    900 

825 

$1500 
1000 

3000  ft.  3-in.  pipe  

300 

750 

1050 

1000  ft.  2-in.  pipe  .      ... 

30 

120 

150 

20,000  ft.  IHn.  pipe  

300 

1,600 

1900 

17,500  ft.  1-in.  and  f-in.  pipe  

130 

770 

900 

Hose,  hose  bibbs,  valves,  elbows,  etc 

500 

Total  

$7000 

The  cost  of  this  plant  is  somewhat  higher  than  would  be 
necessary  elsewhere,  due  to  the  fact  that  very  seldom  would 
both  tank  and  pump  be  required.  The  above  figures,  more- 
over, are  for  a  well-developed  mine  with  distances  somewhat 
well  advanced  from  the  surface.  Here  also,  the  following  of 
a  systematic  plan  from  the  start  for  laying  pipe,  instead  of  pro- 
ceeding hit  or  miss,  would  have  diminished  the  amount  of  pipe 
required  considerably.  Even  installed  at  the  above  cost  and 
on  a  tonnage  of  about  300,000  tons  per  year,  the  entire  plant 
could  be  paid  for  in  one  year  at  cost  of  2a/3c.  per  ton  and  if 
the  cost  were  distributed  over  a  period  of  five  years,  less  than 
y2c.  per  ton  would  suffice,  which  is  certainly  not  prohibitive. 

The  operating  cost  in  1908  was  about  as  follows : 

One  pipe  man,  275  days  at  $2.75  (he  also  sprinkles). .  $  756.25 

One  water  man,  275  days  at  $2.75 756 . 25 

20,000  gal.  of  water  per  day  for  275  days  at  12c.  per 

1000  gal 660.00 

Powerhouse  expense  including  labor  running  pump, 

coal  for  steam,  etc 275 . 00 

6  per  cent  interest  on  investment 420 . 00 

Depreciation  10  per  cent 700 . 00 

Extensions,  repairs,  etc.  (material) 500 . 00 

Total $4,067.50 

The  sum  of  $4067.50  per  year  on  300,000  tons  amounts  to 
lVsc-  Per  ton>  a  mere  trifle  compared  to  benefits  derived  as  will 


502  COAL  MINING  COSTS 

be  quickly  admitted  by  any  company  which  installs  tne  system. 
In  the  above  operating  cost  the  12c.  per  1000  gal.  is  variable 
even  for  this  mine,  and  covers  the  cost  of  bringing  the  water 
to  the  storage  tank.  This  cost  has  at  times  amounted  to  $1.50 
per  1000  gal.,  when  water  was  hauled  40  miles  in  railroad  cars 
and  even  at  this  cost  sprinkling  was  kept  up,  though  its  extent 
was  somewhat  restricted.  Depreciation  at  10  per  cent  is  liberal 
as  there  is  little  wear  on  tank  or  pressure  pump,  and  the  pipe 
will,  under  ordinary  treatment,  last  6  or  8  yr.,  even  where  laid 
on  coal  debris  and  subject  to  the  corrosive  action  of  acids  pro- 
duced by  leaching  of  coal.  Material  for  extensions,  repairs, 
etc.,  will  amount  to  more  than  $500  per  year  when  a  mine  is 
new  and  ground  being  opened  up  fast,  but  during  the  period 
of  retreating  there  will  not  only  be  no  new  material  needed, 
but  pipe  fittings,  etc.,  will  accumulate,  hence  $500  is  placed  as 
an  average.  Power-house  expense  is  the  cost  of  running  the 
pump  at  the  power  house  to  pump  water  from  the  tank  to  the 
mine,  and  includes  attendance  of  engineer,  who  also  attends 
air  compressor,  dynamos,  etc.,  as  well  as  cost  of  coal,  water, 
labor,  etc.,  used  in  generating  steam  to  run  pump. 


INDEX 


Acceleration  in  haulage,  227 
Adhesion     of    steel     and     cast-iron 

wheels  to  the  track,  228 
Adjustable-turret    arc    wall    cutting 

machines,  111 

Africa.     See  "South  Africa" 
Air,  compressed.     See  "  Compressed 

Air" 
Alternating  current: 

Layout  for  mine  using  alternating 

current  cutting  machines,  109 
Anemometer  tests,  444 
Animal  haulage  costs.     See  "Haul- 
age" 
Anthracite,    average   and   bulk   line 

costs  of,  17 
Anthracite : 

Increase  in  costs  after  second  War 

Bonus,  18 

Percentage  of  different  sizes  of,  15 
Prices  fixed  by  the  President,  Aug. 

23,  1917,  14,  19 
Prices    received    for    White    Ash, 

average,  17 

Prices  fixed,  Dec.  31,  1918,  20 
Royalties  on,  1,  16 
U.  S.  Census  Report  on,  for  1909,  1 
Anthracite  Coal  Waste  Commission 
estimates  of  percentage  of  recov- 
ery, 164 
Apportionment    of    costs    to    cover 

future  operation,  31 
Arc  wall  cutting  machines,  111 

B 

Back  haul,  222 

Bearings,  roller,  for  mine  cars,  227. 
301 


Bits  for  mining  machines,  102,  115 

See  also  "Mining  machines" 

Number  required,  116 

Sharpening,  115 

Sullivan  Machinery  Co.,  "Dread- 
naught,"  115 

Tempering,  116 
Blasting,  125 

Dynamite,  128,  177 

Guncotton,  130 

Missed  shots,  138 

Nitroglycerin,  129 

Shaft  sinking,  177 
Bonding  rails,  267 

Compressed  terminal  bonds,  267 

Copper  wire  resistance,  272 

Efficiency  of,  271 

Losses  in  bonding,  268,  272,  275 

Rail  to  copper  ratio,  273 

Solid  terminal  bonds,  270 

Wire  rope  for,  276 
Brattices,  488,  495 
Breakage  of  coal.     See  "Screenings" 
Buying  mine  locomotives,  232 


Candlepower,  456 

Portable  lamps,  458 
Capacity : 

Mining  and  loading  machines.    See 
under  that  title 

Mining  machines.     See  under  that 

title 
Capital  invested: 

Anthracite  mines,  2 

Bituminous  mines,  3 

Illinois  and  Indiana  mines,  5 

Westphalia  (Germany)  mines,  33 
Cars,  mine,  299 


503 


504 


INDEX 


Cars,  Capacity  of,  299 

Frictional  resistance  of,  224,  227, 

230 

Oiling,  302 
Repair  costs,  300 
Roller  bearings  for,  301,  303 
Steel  cars,  299,  300 
Stretcher  cars,  303 
Tare  and  weight,  300 
Wheels  for,  303 
Cartridges,  hydraulic,  131,  134 
Advantages  of,  135 
Capacity,  135 
Commercial  value  of,  133 
Percentage  of  large  coal  obtained 

with,  132 
Test  of,    at  the  Hulton   Colliery 

(England),  132 
Cement     gun     for     timbering.     See 

"Timbering" 

Census  reports,  U.  S.  for  1909,  pro- 
duction, costs,  salaries,  etc.,  1 
Charges.     See    subject    as    "Deple- 
tion,"   "Depreciation,"    "In- 
terest," etc. 

Coke,  maximum  capacity  of  drawing 
ovens  at  the  operations  of  the 
U.  S.  Coal  &  Coke  Co.,  147 
Colorado : 

Systems  of  mining  and  percentages 

of  recovery  in,  166 
Detailed    cost   of   sinking    570-ft. 

shaft,  198 
Comparative  cost  of  different  systems 

of  haulage.     See  "Haulage" 
Compressed  air: 

Locomotive    haulage     costs.     See 

"Haulage" 
Mining  machines.     See  under  that 

title 

Underground  compressor  for  driv- 
ing conveyor,  95 
Compressed  terminal  bonds.        See 

"Bonding" 
Compressors  (air) : 
for  shaft  sinking,  186 
Single-  and   two-stage   compared, 
186 


Concentration  method: 
Connellsville  district,  81 
Gary,  W.  Va.,  48 
Concrete : 

Shaft     linings.     See    under     that 

title 

Timbers.     See  under  that  title 
Connellsville  district : 

Frick  Co.  system  of  mining,  82 
Method   of   laying   out   in    90-ft. 

blocks,  87 

Method  of  working  at  the  Conti- 
nental mine  No.  2,  86 
Systematizing  work  in  rooms,  85 
Systems  of  working,  80 
Conservation : 

Anthracite  Coal  Waste  Commission 

on,  164 
Complete    extraction   required   at 

German  mines,  33 
Economic  aspects  of,  164 
Factors  governing  percentages  of 

recovery  in  various  fields,   168, 

170 
Percentages    of    recovery    at    the 

mines  of  the  Pocahontas  C.  &  C. 

Co.,  79 
Summary    of   results    in    different 

fields,  165 

Use  of  longwall  in,  173 
Continuous  panel  system  of  working 

at  Gary,  W.  Va.,  52 
Contract  form  for  shaft  sinking,  217 
Conveyor  system  of  mining,  93,  95 
Copper  wire,  resistance  of,  in  bond- 
ing, 272 
Costs: 

See  under  various  heads 
Apportionment  of,  to  cover  future 

operations,  31 
Central  Pennsylvania,  26 
Illinois  No.  6  District,  24 
Indiana,  24 

Increases  in,  from  1916  to  1920,  27 
Increase  in,   due    to    intermittent 

work,  161 
Influence  of  thickness  of  seam  on, 

22 


INDEX 


505 


Costs: 

Labor.     See  under  that  title 
Middlewestern  and  German  mines 

compared,  31 
Ohio  No.  8  district,  25 
Pocahontas  field,  25 
Southwestern  Pennsylvania,  23 
Cottonwood  Coal  Co.,  shaft  sinking 

report,  form  of,  215 
Crozer  Land  Association  system  of 

mining,  72 
Curtain  wall  for  shafts.     See  "Shaft 

linings" 
Curves,  track.     See  "Track" 

D 

Daymen,  139 

Average  number  required  at  454 

mines,  140 
Effects  of  the  thickness  of  coal  on 

number  required,  142 
Efficiency  of,  60 
Production  per  dayman,  141,  142, 

145 
Variations    in,    with    capacity    of 

mine,  142 
Depletion     charges     at     anthracite 

mines,  20 
Depreciation : 

Charges  at  anthracite  mines,  20 
Depth  of  undercutting,  108 
Developed   coal  property  compared 

with  an  undeveloped,  28 
Development  work: 
Rate  of,  53,  95 
Shaft    sinking.     See    under    that 

title 

Direct  current  compared  with  alter- 
nating for  operating  machines, 
109 
Dominion  Coal  Co.  loading  machines, 

118 

Doors,  underground,  492 
Drainage.     See  "Pumping" 
Drawbar,  determining  height  of,  229 
Drawbar-pull    of    locomotives,    222, 

224,  231 
Drawing  pillars.     See  "Pillars" 


Drills,  196 

Churn,  180 

for  soft  material,  180,  188 

Tests  of,  on  the  Transvaal,  433 

Steel  for,  437 
Drilling : 

Churn,  180 

for  shaft  sinking,  176,  177 

hand,  177,  189,  215 

in  soft  material,  180,  188 

records  made  in  the  Transvaal,  433 
Dynamite : 

See  also  "Blasting,"  "Explosives" 

Fumeless,  177 

Gelatine,  177 

E 

Efficiency: 

See  also  "Loading  machines" 
Determining  maximum,  66 
Hand  and  machine  loading  com- 
pared, 117,  120 
Mining  machines,  102 
of  daymen,  materials  and  equip- 
ment, 60 
of  various   companies  in   Central 

Pennsylvania,  145 
Outside  handling  systems,  150 
Per  capita  production  and  percent- 
age of  machine  mined  coal,  98 
Production  per  employee,  98,  148, 

152,  154 

Shoveling  capacity  of  miners,  156 
Electric  lamps.     See  "Lighting" 
Electrical    shotfiring.      See     "Shot- 
firing" 
Elizabeth  tunnel  on  the  Los  Angeles 

Aqueduct,  422 

Employees.     See    "Work,"    "Load- 
ing," "Efficiency,"  etc. 
Endless  rope  haulage,  307 
Engineering,  effects  of  in   obtaining 

a  maximum  recovery,  171 
Engines     for     driving     underground 

conveyors,  95 

English  shaft  sinking  methods,  203 
Equipment   for   shaft    sinking.     See 
"Shaft  sinking  equipment" 


506 


INDEX 


European  shaft  sinking  methods : 
Belgium,  203 
English,  203 

Evans  scraper  loading  apparatus,  119 
Excavating  for  shafts.     See  ''Shaft 

sinking" 
Exhaustion    of    coal    reserves.     See 

"  Conservation  " 
Exploders  used  by  the  Utah  Fuel  Co., 

137 
Explosives: 

See  also  "Blasting" 

Exploders  "Reliable,"  137 

for  tunneling,  440 

Giant  powder  used  by  the  Utah 

Fuel  Co.,  137 
Permissible,  130 

Production  and  amount  used  in  the 
U.  S.,  130 


Fans,  ventilating,  443 

Anemometer  tests  of,  444 

Efficiency  of,  443 

Manometric  efficiency  of,  444 

Mechanical  efficiency  of,  444,  452 

Power  required  for,  446 

Specifications  for,  448 

Volumetric  efficiency  of,  444 

Water  gage  requirements,  450 
Fluctuations  in  tonnage  as  effecting 

costs,  22,  32,  34,  39,  161,  162 
Forces  in  blasting,  127 
Frick  Co.: 

Shafts  at  Brownsville,  Pa.,  192 

Systems  of  mining,  82 
Frictional  resistance.     See   "Resist- 
ance" 
Frogs  and  switches,  291 

Cost  of  laying,  293 

Spacing  of  ties,  294 
Fuel  Administration: 

Costs    on    anthracite,    December, 
1917,  to  May,  1918,  20 

Costs,   prices  fixed  and  tonnages, 
August  12,  1918,  5,  8 

Costs,   prices  fixed  and  tonnages, 
full  year,  1918,  5 


Fuel  Administration: 

Costs,  prices  fixed  and  tonnages, 

previous  to  November,  1917,  7 
Prices  fixed  on  anthracite,  Decem- 
ber 31,  1918,  20 

Reported  costs,  prices  fixed  and 
tonnages  for  the  full  year,  1918, 
13 

Fuel  and  power  costs  at  mines  com- 
puted   by    the    U.    S.    Census 
Bureau  for  1909,  1 
Future  worth  of  coal  properties,  29,  51 


Gary,    W.    Va.,    percentages   of   re- 
covery at,  80 
Gathering  locomotives,  251 

See  also  "Locomotives" 
Georges  Creek  field: 

Systems  of  mining,  39 

Percentages  of  recovery,  168 
Germany : 

Costs  of  mining  and  distribution  of 
revenue,  31 

Production  per  employee,  152 

Shapes  of  shafts,  185 
Grade,  track: 

Effect  of,  on  locomotives,  229 

Weight  transfer  due  to,  229 

See  also  "Track" 
Gravity  planes,  307 
Great  Britain: 

See  also  "England" 

Production  per  employee  compared 
with  the  U.  S.,  154 

Shaft  sinking  methods,  203 
Grounding  losses.     See  "Line  costs 

and  losses" 
Guncotton,  130 


Hand  drilling.    See  "Drilling" 
Hand  shoveling  compared  with  ma- 
chine loading,  117,  120 
Haul,  increasing  entry  driven  to  ob- 
tain shortest,  220 
Back  haul,  222 


INDEX 


507 


Haulage,  219 

Acceleration,  227 

Animal,  compressed  air  and  electric 
haulage  costs  compared,  318 

Animal  haulage  costs,  318,  319, 331, 
334,  346 

Capacity  of  locomotives,  224 

Comparative  cost  of  different  sys- 
tems, 314 

Compressed  air  locomotive  costs, 
317,  319,  331,  362 

Compressed  air  single-  and  two- 
stage  locomotives,  327 

Compressed  air  and  animal  haulage 
costs  compared,  331 

Compressed  air  and  electric  haul- 
age costs  compared,  362 

Electric  locomotive  costs,  315,  319, 
334,  357,  362 

Electric    locomotiv-e    and    animal 
haulage  costs  compared,  334 

Gasoline  locomotive  costs,  316,  346 

Gasoline    locomotive    and    animal 
haulage  costs  compared,  346 

Gathering  methods,  58,  156 

Gravity  planes,  307 

Number  of  cars  per  trip,  286 

Preliminary  considerations,  219 

Rope,  306 

Rope  haulage  costs,  317 

Storage  battery  locomotive  costs, 
316,  345;  357 

Storage  battery  and  trolley  loco- 
motive costs,  357 

Horse  haulage  costs .     See  ' '  Haulage ' ' 
Horsepower  of  mine  locomotives,  224 
Horsepower  rating,  231 
Hydraulic  cartridges.    See  also  "Car- 
tridges," 131 


Illinois: 

Costs  in  the  No.  6  District,  1916  to 

1920,  24 

Efficiency  of  outside  plants,  150 
Increase  in  costs,  1916  to  1920,  27 
Systems  of  mining  and  percentages 
of  recovery,  166 


Illinois: 

U.  S.  Census  report  of  wages,  sala- 
ries, royalties,  etc.,  for  1909,  5 
Inclined     shafts     and     slopes.     See 

" Shafts"  and  " Shaft  sinking" 
Increases  in  costs,  1916  to  1920,  27 
Indiana : 

Costs  for  years  1916  to  1920,  24 
Increase  in  costs  1916  to  1920,  27 
U.  S.  Census  report  of  wages,  sala- 
ries, royalties,  etc.,  in  1909,  5 
Ingersoll-Rand  Mining  and  loading 

machine,  122 
Inside  men.    See  "Daymen,"  "Labor 

costs,"  "Work" 
Interest  charges,  increase  of,  due  to 

intermittent  work,  163 
Intermittent  work: 

at  mines  compared  with  other  in- 
dustries, 162 

Comparison  of  shifts  worked  at 
Middlewestern  and  Westphalia 
mines,  34 

Increase  in  costs  due  to,  161 
Influence  of  on  costs,  22,  39 
Losses  to  miners  due  to,  32,  157 
Peabody,  F.  S.,  on,  162 


Jeffrey-Drennan  adjustable-turret 
cutting  machine,  111 

K 

Kentucky,   machine  mined  coal  in, 
98 


Labor,  effects  of  union  rules  in  obtain- 
ing maximum  recovery,  172 
Labor  costs: 

See  also  "Daymen" 

Average  proportion  of  cost,  139 

Central     Pennsylvania,     1916    to 

1920,  26 
Illinois  No.  6  District,  1916  to  1920, 

24 
Indiana,  1916  to  1920,  24 


508 


INDEX 


Labor  costs : 

Labor  costs  at  454  mines,  140 
Maximum  labor  effort  as  developed 

in    tests    by    the  U.  S.  Coal   & 

Coke  Co.,  145 
Ohio  No.  8  District,  1916  to  1920, 

25 
on  anthracite,  December,  1917,  to 

May,  1918,  20 

on  anthracite,  1913  to  1916,  23 
Pocahontas  field,  1916  to  1920,  23 
Production  per  dayman,  141,  142, 

145 

Production  per  employee,  148 
Ratio  to  total,  21 
Southwestern  Pennsylvania,    1916 

to  1920,  23 
Variations    in,    with    capacity    of 

mine,  142 
Lamps,  miners: 

See  also  "Lighting" 

Acetylene  and  oil,  470 

Cables  for,  467 

Charges  to  miners  for,  469 

Electric  bulbs  for,  466 

Maintenance  costs,  461,  470 

Manlite,  460 

Oil  lamps,  471 

Plant   cost   for   installing   electric 

lamps,  471 
Lighting,  mine,  456 

Acetylene  and  oil  lamps,  470 
Bureau  of  Mines  standards,  458 
Charges  to  miners  for,  469 
Lumens,  456 
Measurement  of,  456 
Oil  lamps,  471 
Portable  lamps,  459 
Line  costs  and  losses,  246,  264 
Cost  of  feed  wire,  265 
Cost  of  line  construction,  267 
Drop  in  voltage,  264 
Grounding  losses,  266 
Linings,  shaft.    See  "  Shaft  linings  " 
Loading : 

See  also  "Loading  machines" 
Hand  and  machine  loading  com- 
pared, 117 


Loading: 

Maximum  effort  in  hand  loading 
developed  at  tests  by  the  U.  S. 
Coal  &  Coke  Co.,  145 

Number  of  tons  loaded  per  shift  by 
miners  in  the  Connellsville  Dis- 
trict, 85 

Shoveling  capacity,  156 
Loading  and  mining  machines,  122 

See  also  "Mining  and  loading  ma- 
chines" 
Loading  machines,  116 

Capacity  of,  118 

Evans  scraper,  119 

Hand  shoveling  compared  with, 
117,  120 

Jeffrey  loader,  117,  118 

Layout  of  mine  for  machine  load- 
ing, 117 

Meyers- Whaley  loader,  120 

Power  required  for,  120 

Westmoreland  loader,  120 
Locomotives: 

Adhesion  of  cast-iron  and  steel 
wheels  on,  228 

Ampere  rating  of,  233 

Buying,  232 

Checking  the  work  of,  253 

Commutating  and  non-commutat- 
ing  pole  motors,  256 

Comparison  of  one  hour  and  all  day 
rating,  234 

Computing  size  of,  223,  236 

Costs  of,  252 

Curves  for  a  40-hp.  locomotive, 
235 

Curves  for  a  50-hp.  locomotive, 
240 

Effect  of  long  feed  lines  on,  263 

Effective  wattage,  224 

Haulage  capacity  of,  224 

Haulage  costs  with.  See  "Haul- 
age" 

Horsepower  of,  224,  233,  245 

Height  limits  of,  248 

Large  size,  250 

Life  of,  250 

Motor  losses  on,  253 


INDEX 


509 


Locomotives : 

Number  and  size  required,  233,  249 

One  hour  rating  of,  233 

Power  costs  for,  256,  260 

Resistance,  227 

Roundtrip  performance  of,  261 

Speeds,  224 

Tractive   effort   drawbar-pull   and 

rating  of,  222,  224,  235,  245 
Tractive  effort  per  inch  of  height, 

248 

Longwall,  use  of,  in  conservation,  173 
Loss: 

Conditions  where  operations  may 
be  conducted  at,  and  apparent, 
33 

Mining  unprofitable  seams,  93 
Lump  coal,  increase  of,  with  mining 
machines,  112 


M 

Machines : 

Loading.     See      "  Loading     ma- 
chines" 

Mining.    See  "Mining  machines" 
Mining  and  loading.    See  "  Mining 

and  loading  machines" 
Management,  effect  of,  in  obtaining 

maximum  recovery,  171 
Mexico,  details  of  sinking  a  679-ft. 

shaft  in,  198 

Michigan,  systems  of  mining  and  per- 
centages of  recovery  in,  166 
Middlewestern  mines: 

Comparison  of  costs  and  distribu- 
tion of  revenue  compared  with 
German  mines,  31 
Comparison  of  prices  realized 
with  those  of  Westphalia,  Ger- 
many, 34 

Progress  in  shaft  sinking,  213 
Miners : 
See  also  "Loading  machines"  and 

"  Daymen" 

Earnings  of.     See  "Wages" 
Hand  and  machine  loading  com- 
pared, 117,  120 


Miners: 

Maximum  loadings  of,  in  tests  by 
the  U.  S.  Coal  &  Coke  Co.,  145 

Number  of  tons  loaded  per  shift  in 
the  Connellsville  region,  85 

Shoveling  capacity  of,  156 

Tons  mined  per  employee,  98,  148, 

152,  154 
Mining  and  loading  machines,  122 

Capacity  of,  123 

Ingersoll-Rand  type,  122 

Jeffrey  type,  123 

O'Toole  type,  124 

Saving  of  timbering  with  the  use  of 

machines,  124 
Mining  machines: 

Alternating  current  for,  109 

Arc  wall  cutters,  111 

Bits  for.    See  also  "Bits,"  102, 115 

Bonus  system  in  operating,  107 

Buying  machines,  101 

Capacity  of,  107,  111,  112 

Care  of,  102,  115 

Charges  against,  100 

Considerations  affecting  their  ad- 
aptability, 99 

Cost  of,  99 

Cutting  machines,  97 

Economies  of,  97 

Hand  and  machine  operating  condi- 
tions compared,  101,  104,  105 

Installation  costs,  103 

Jeffrey-Drennan  adjustable-turret 
machines,  111 

Labor  and  mechanical  power  com- 
pared, 96 

Loading  machines.  See  "Loading 
machines" 

Lump  coal  proportion  with  ma- 
chines, 112 

Maintenance  costs,  107 

Maintenance  costs  for  alternating 
and  direct  current  installations, 
111 

Obtaining  efficiency  from,  102 

Operating  costs,  103 

Post  punching  machines,  111 

Power  for,  97,  103,  109 


r 


510 


INDEX 


Mining  machines: 

Proportion  of  coal  cut  by,  97,  99 

Repair  costs,  112,  115 

Repair  reports,  113 

Shortwall  used  in  the  Connellsville 

district,  81 
Small  capacity  mines,  high  costs  at, 

142 

Sullivan  Machinery  Co.   "Dread- 
naught"  chain,  115 
Supplies  for,  accounting  of,  114 
Undercutting,  depth  of,  66,  108 
Mining,  systems,  39 

Concentration    method    used    at 

Gary,  W.  Va.,  48 
Connellsville  district,  80 
Continental  mine  No.  2  (Connells- 
ville district),  86 
Continuous  panel,  52 
Conveyor  system,  93 
Crozer  Land  Association,  72 
Driving  rooms .     See  ' '  Rooms ' ' 
Frick  Co.  methods,  82 
Georges  Creek  field,  39 
in  various  fields  with  special  refer- 
ence to  percentages  of  recovery, 
165 
Laying  out  in  90-ft.  square  blocks, 

87 

Layout  for  machine  loading,  117 
Layout    for    alternating     current 

mine,  109 

Pocahontas  field,  72 
Square  or  rectangular  panel,  52 
Upland  Coal  &  Coke  Co.,  72 
Working  thick  soft  seams,  45 
Missed  shots.     See  "Blasting" 
Mule  haulage  costs.     See  "Haulage" 
Mules: 

Cost  and  care  of,  338 
Depreciation  of,  340 
Work  of,  340 
Meyers- Whaley  loading  machine,  120 

N 
New    York    Aqueduct,    progress    in 

shaft  sinking  at,  213 
Nitroglycerin,  129 


Ohio: 

Costs  in  the  No.  8  District,  1916  to 

1920,  25 

Increase  in  costs,  1916  to  1920,  27 
Percentage  of  recovery  in,  169 
Oil: 

Cost  of  for  mine  cars,  302 
Illuminating,  471 
Outside       men.       See    "Daymen," 

"Labor  costs,"  "Work" 
Outside  plants,  the  most  efficient,  150 
Outputs : 

Comparison   of   shifts  worked   at 

Middlewestern  and  Westphalia 

(Germany)  mines,  34 

Estimating,  from  a  mine  section,  90 

Influence    of    fluctuations    in,    on 

costs,  22,  32,  39,  161,  164 
Overcasts,  488,  492 


Pacific  Coast  Coal  Co.,  post  punchers 

at,  111 

Panel  system  of  mining: 
Continuous,  52 
Square  or  rectangular,  52 
Pennsylvania,  U.  S.  Census  report  on 

anthracite,  1 
Pennsylvania  bituminous: 

Costs  in  the  Central  District,  1916 

to  1920,  26 
Costs    in    Southwestern    District, 

1916  to  1920,  23 
Increase  in  Central  District  costs, 

1916  to  1920,  27 
Increase  in  Southwestern  District 

costs,  1916  to  1920,  27 
Systems  of  mining  and  percentages 

of  recovery,  167 

Percentages  of  recovery.    See  "Con- 
servation" and  "Drawing  pillars" 
Permissible     explosives.       See     also 

"Explosives" 
Pillars: 

Barrier,  56,  60,  69 
Size  of.     See  various  systems  of 
mining 


INDEX 


511 


Pillars,  drawing: 

See  also  various  systems  of  mining 
Concentration     method     used    at 

Gary,  W.  Va.,  49 
Georges  Creek  field,  47 
Maintenance  of  breakline,  77 
Methods  in  the  Connellsville  re- 
gion, 88 

Percentage  of  recovery  at  the  mines 
of  the  Pocahontas  Coal  &  Coke 
Co.,  79 
Percentages  of  recovery  at  Gary, 

W.  Va.,  80 

Pocahontas  field,  74,  78 
Use  of  mining  machines  in,  77 
Pocahontas  Coal  &  Coke  Co.,   sys- 
tem of  mining,  72 
Pocahontas  field: 

Methods  of  drawing  pillars,  74 
Methods  of  working  in,  72 
Shaft  water  in,  187 
Portable  lamps .     See  ' '  Lighting ' ' 
Post  punching  machines,  111 
Power : 

Comparison    of    alternating    and 
direct  current  for  operating  ma- 
chines, 109 
Costs  of  for  operating  locomotives. 

See  "Locomotives" 
Plant  costs  by  various  units,  261 
Power  and  fuel,  costs  of  reported  by 
the  U.S.  Census  Bureau  of  1909, 1 
Present   and   future   worth   of    coal 

properties,  29 
Prices: 
Anthracite,  fixed  by  the  president 

in  1917,  14,  19 
Comparison  of,  at  Middlewestern 

and  Westphalia  mines,  34 
Fixed  by  the  Fuel  Administration. 

See  under  that  head 
Method  of  fixing  used  by  the  West- 

phalian  Coal  Syndicate,  32 
Production: 

Comparison     of,     at     Westphalia 
(Germany)    and   Middlewestern 
mines,  33 
Influence  of,  on  costs,  22,  39 


Profits: 

as  affecting  conservation,  172 
Increase  of,  with  increased  output, 

164 

U.  S.  Census  report  on,  for  1909,  4 
Profit    and    loss,    conditions    where 
operations  may  be  conducted  at 
an  apparent  loss,  33 
Pumping.    See  "  Shaft  sinking,  pump- 
ing" 
Punching  machines : 

See  also   "Mining  machines" 
Cutting    and    punching    machines 

compared,  106 

Post  punching  machines,  111 
Pulling  pillars.     See  "Pillars" 

R 

Rails,  track,  287 
Durability,  289 

Electrical  resistance  of.    See  "Lo- 
comotives" 
Frogs.     See  "Frogs" 
Minimum  weight,  287 
Purchasing,  287 
Seconds,  290 
Stiffness,  288 
Strength,  288 
Recovery  of  coal : 

See  also  "Conservation" 
Anthracite  coal  waste,  Commission 

on,  164 
Factors    governing    recovery    in, 

various  fields,  168,  170 
Percentage  of,  in  various  fields,  165 
Percentage    of,    at   mines    of   the 
Pocahontas  Coal  &   Coke  Co., 
79 

Records  of,  79 
Use  of  longwall  to  effect  maximum, 

173 

Refuge  chambers,  498 
Report  form  for  shaft  sinking,  215 
Reports  for  repair  costs  to  mining 

machines,  113 

Resistance,  track  and  grade,  223,  231 
of  cars,  224,  227,  230 
Locomotives,  227 


512 


INDEX 


Resultant  forces  in  blasting,  127 
Revenue  distribution  at  German  and 

Westphalia  mines,  31 
Robbing  pillars.     See  "  Pillars" 
Rooms : 

See  also  various  systems  of  mining 

Depth  and  number  of,  78 

Methods    of    driving     commonly 
used,  49 

Number  of  men  in,  50 

Placing  track  in,  49 

Room  space  per  miner,  67 

Safety  of,  51 

Systematizing  wor.i  in,  85 
Roller  bearings .     See  ' '  Bearings ' ' 
Rope  haulage,  306 
Rope,  wire,  309 

Lubrication,  310 
Royalties : 

on  anthracite,  16,  20 

U.  S.  Census  report  on,  for  anthra- 
cite mines  in  1909,  1 

U.  S.  Census  report  on,  for  bitu- 
minous mines  in  1909,  3 

U.  S.  Census  report  on,  for  Indiana 
and  Illinois  mines  in  1909,  5 


S 
Salaries: 

U.  S.  Census  report  on,  for  anthra- 
cite mines  in  1909,  1 
U.  S.  Census  report  on,  for  bitu- 
minous mines  in  1909,  3 
U.  S.  Census  report  on,  for  Indiana 

and  Illinois  mines  in  1909,  5 
Charges  for,  at  anthracite  mines, 
December,  1917,  to  May,  1918, 
20 

Scales,  wage.     See  "Wage  scale" 
Scraper  loading  machine,  Evans,  119 
Screenings : 

Breakage  with  hydraulic  cartridge, 

132 

Proportion  of,  with  machine  min- 
ing, 112 

Seams,  costs  of  mining  different  thick- 
nesses of,  22,  91 


Shafts: 

Circular,  179,  181,  185,  208 

Elliptical,  178,  182,  183,  185,  192 

Hydrostatic  pressure  in,  181 

Inclined,  comparison  of  cost  with 
vertical,  189 

Preliminary  considerations  for,  176 

Prospect,  203 

Quadrilateral,  182,  185 

Rectangular,  177,  208 

Shapes,  177,  208 

Weep  holes  in,  181 
Shaft  linings,  204 

Brick,  203 

Comparison  of  different  shapes,  182 

Concrete,  181,  192,  195,  203,  208 

Concrete  blocks,  207 

Concrete  shaft  for  the  River  Coal 
Co.,  206 

Curtain  wall,  193 

Grouting,  181 

Hydrostatic  pressure  against,  181 

Timber,  201,  208 

Weep  holes  in,  181 
Shaft  sinking,  176 

Blasting,  177,  193 

Caisson,  192 

Circular,  179,  208 

Contract  form  for,  217 

Drainage.     See  "Pumping" 

Drilling,  176,  188 

Elliptical,  178 

Equipment  for.     See  "Shaft  sink- 
ing equipment" 

Excavation,  182,  184,  191, 194,  199, 
205 

Grouting,  181 

Hand  drilling,  177,  189,  215 

Hydrostatic  pressure  in,  181 

in  quicksand,  191 

in  soft  material,  180,  188 

Inclined,  comparison  of  cost  with 
vertical,  189 

Operation,  176,  193,  204 

Report  form  for,  215 

South  Africa,  180 

through  surface  material,  191 

Wages  in,  180 


INDEX 


513 


Shaft  sinking  costs,  188,  191,  213 

Clay  and  gravel,  191 

Comparison  of  circular  and  rectan- 
gular, 208 

Costs  in  Michigan,  200 

in  Great  Britain,  203 

in  quicksand,  191 

in  soft  material,  188 

Inclined    and    vertical    compared, 

189,  200 
Nokomis  Coal  Co.  (Illinois),  188 

per  cubic  foot  and  cubic  yard,  191 

Power  costs,  197 

Prospect,  203 

Sinking  570-ft.  shaft  in  Colorado, 
198 

Sinking  679-ft.  shaft  in  Mexico,  198 
Shaft  sinking  equipment,    186,    196, 
202,  211 

Boilers,  187 

Compressed  air  plant  for,  186,  196 

Dump  cars,  196 

for  sinking  a  shaft  in  Michigan,  200 

for  sinking  a  prospect  shaft,  203 

Headframe,  187 

Hoist,  187 

Plant  for  sinking  a  single  500-ft. 
shaft,  187 

Pumps,  187 

Single-  and  two-stage  compressors 

compared,  186 

Shaft  sinking,  progress,  194,  197,  205, 
213,  209 

Computed  progress  in  Michigan, 
200 

Effects  of  large  volumes  of  water 
on,  187 

in  Great  Britain,  203 

Middlewestern,  213 

New  York  Aqueduct,  213 

Report  forms  for,  215 

Various  American  and  South  Afri- 
can shafts,  212 
Shaft  sinking,  pumping,  187 
Shifts  worked: 

See  also  "Fluctuations  in  tonnage," 
"Intermittent  work,"  "Out- 
puts" 


Shifts  worked: 

at  mines,  compared  with  other  in- 
dustries, 162 
Comparison  of,  at  Middlewestern 

and  Westphalia  mines,  34 
in  shaft  sinking,  180 
Shooting  coal.    See  "Blasting" 
Shot  firing,  electrical: 
See  also  "Blasting" 
Cost  of  installing,  138 
Power  required  for,  137 
Safety  of,  139 
Wiring  for,  136 

Shoveling  capacity.  See  "Effici- 
ency," "  Loading, "  "  Miners, ' ' 
"Work" 

Sinking  shafts.  See  "Shaft  sinking" 
Sizes  of  anthracite,  percentage  of,  15 
South  Africa: 

Shaft  sinking  practice,  211 
Wltwatersrand,  213 
Speed  of  development  work,  53,  95 
Speed  of  shaft  sinking.     See  "Shaft 

sinking,  progress" 

Square  panel  used  at  Gary,  W.  Va.,  52 
Stables,  underground,  475 
Steel  mine  cars.    See  "Cars,  mine" 
Steel  for  timbering.     See  "Timber- 
ing" 

Stoppings,  488 
Storage    battery    locomotives.      See 

"Haulage,"  "Locomotives" 
Sprinkling  costs,  499 
Stretcher  car,  303 
Switches,    track.      See    "Frogs   and 

switches" 
Systems  of  mining: 

Concentration    method    used    at 

Gary,  W.  Va.,  48 
Connellsville  district,  80 
Continental  Mine  No.  2  (Connells- 
ville district),  86 
Continuous  panel,  52,  60 
Crozer  Land  Association,  72 
Driving  rooms.     See  "Rooms" 
for  working  thick  seams,  45 
Frick  Co.,  82 
Georges  Creek  field,  39 


514 


INDEX 


Systems  of  mining: 

Laying  out  in  90-ft.  square  blocks, 
87 

Layout  for  an  alternating  current 
mine,  109 

Layout  for  machine  loading,  117 

Longwall  used  to  affect  conserva- 
tion, 173 

Pocahontas  field,  72 

Square  or  rectangular  panel,  52,  60 

Upland  Coal  &  Coke  Co.,  72 


Taxes,  computing  returns  for,  28 
Thickness  of  seam,  influence  on  costs, 

22,  91 
Ties,  track: 

Comparison  of  steel  and  wood,  295 

Life  of  steel  ties,  296 

Spacing  of,  for  switches,  294 
Timber: 

Computing  size  of,  370 

Concrete  for  timbering,  393 

Costs  and  amounts  used  in  the  U. 
S.,  367 

Framing  equipment  and  costs,  376 

How  to  buy,  365 

Kinds  used,  369 

Knots,  effects  of,  on  strength,  374 

Preservatives,  379,  382 

Reclaiming,  404 

Round  and  sawed,  369 

Seasoning,  380 
Timbering : 

Cement  gun  used  for,  402 

Computing  sizes  for,  370 

Costs,  365 

Linings   for   shafts.      See    " Shaft 
linings" 

Maintenance    cost    of    wood    and 
steel,  390 

Rails  for,  374 

Savings  in  with  mining  and  loading 

machines,  124 

Time  required  to  reach  a  certain  out- 
put, 53,  95 
Tonnage,  influence  of,  on  costs,  22 


Track: 

Costs,  278,  294 

Cost  of  grade  revisions,  283 

Curvature  of,  in  relation  to  wheel 
base,  232 

Curves.  285 

Frogs.     See  under  that  title 

Grade  revision,  280 

Grades  of,  278,  280 

Methods  of  placing,  in  rooms,  49 

Rails.     See  under  that  title 

Resistance,  223,  279,  298 

Ties.     See  under  that  title 
Tractive  effort  of  mine  locomotives, 

222,  231 
Tunneling  costs,  406,  419 

American  and  foreign  records  com- 
pared, 406 

Bonus  system  applied  to,  412 

Elizabeth  tunnel  on  the  Los  An- 
geles aqueduct,  422 

Explosives  for,  440 

Gunnison   and  Simplon  tunneling 
records,  406 

Utah  Fuel  Co.  methods,  413 
Turnouts.     See  " Frogs  and  switches" 

U 

Undercutting,  depth  of,  108 

See  also  "Machine  mining" 
U.  S.  Census  report  on  production, 

costs,  wages,  salaries,  1 
Upland  Coal  &  Coke  Co.,  system  of 

mining,  72 


Value: 

Developed  and  undeveloped  prop- 
erties, 28 

Present  and  future  worth,  29,  51 
Ventilation : 

See  also  various  systems  of  mining 

Air  required,  453 

Canvas  tubing  for,  455 

Connellsville  district,  81 

Consolidation  Coal  Co. 's  methods, 
450 

Costs,  443 


INDEX 


515 


Ventilation: 

Fans.     See  under  that  title 
Water  gage  readings,  450 

W 
Wage,  ratio  to  value  of  product,  5,  21, 

139 
Wage  scales: 

Hocking  district,  1898  to  1921,  158 
How  fixed,  22 
Wages: 

Causes  for  inadequacy,  155 

Inequalities  of,  155 

Mine   car   supply,    effects   of,    on 

wages,  156 
Miners,  153 
on  shaft  sinking,  180 
U.  S.  Census  report  on,  for  anthra- 
cite mines  in  1909,  1 
U.  S.  Census  report  on,  for  bitu- 
minous mines  in  1909,  3 
U.  S.  Census  report  on,  for  Indiana 

and  Illinois  mines  in  1909,  5 
Westmoreland  coal  loading  machine, 

119 

Westphalian    Coal   Syndicate    (Ger- 
many) : 
Costs  and  revenue  compared  with 

Middlewestern  mines,  31 
Increase    in   production,    1850   to 
1907,  33 


Westphalian    Coal  Syndicate    (Ger- 
many) : 
Investment  per  ton  of  production, 

33 

Price  fixing  at,  32 
Prices     compared     with     Middle- 
western  mines,  34 
WTheels,  mine  car,  303 
Wheels,    adhesive    characteristics   of 
cast  iron  and  steel  for  mine  loco- 
motives, 228 

Wheel  base  and  radius  of  curve,  232 
Wire  rope,  309 

Lubrication  of,  310 
Work: 

See     also     "Loading     machines," 

"  Day  men" 

Hand  and  machine  loading  com- 
pared, 117,  120 
Intermittent  work  and  its  effect  on 

costs,  161 

Maximum  labor  effort  developed  in 
tests  by  the  U.  S.  Coal  and  Coke 
Co.,  145 
Number  of  tons  loaded  per  shift  in 

the  Connellsville  district,  85 
Shoveling  capacity,  156 
Tons  mined  per  employee,  98,  148, 

153 

World,  production  per  employee  in 
various  fields  of,  152 


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